Search Documents

By Andrew Briggs, Denis Kidby

VITROKELER cyanide recovery technology has been successfully piloted at the Gabanintha and Cosmo-Howley mines by Nedpac Engineering Pty. Ltd. The pilot campaigns were designed to confirm technical and economic feasibility of the recovery process including the production of base metal concentrates and return of re-useable cyanide to the leach circuit. The capital and operating costs of full-scale plants is now able to be estimated to within ±20% on the basis of hydrometallurgical data. Firm quotes are subject to on-site test-work, and any special requirements or site peculiarities.

Jan 1, 1990

By S. Luning

In the nineties the gold sector in Burkina became subject to internationally prescribed liberalization policies. The national state is at present only one of many players. Who are these players, and how are they involved with one another in organizing goldmining activities and prospects in Burkina? One mining site serves as a case to study the scope of players and the types of authorities they build upon in governing gold sites. What conclusions can be drawn in regard to the present position of the Burkinabe state within this field of governing authorities?

Jan 1, 2007

By H. Knorre, A. Griffiths, J. Loroesch

Introduction Hydrogen peroxide is known to the mining industry as a powerful oxidant, able to attack and to destroy cyanide, the most important leaching reagent (Knorre and Griffiths, 1984; Griffiths, 1988; Griffiths, 1988). CN- + H202 => OCN- + H2O The most recent development, however, is the use of H202 as an oxidant in the cyanidation of precious metals. This application may appear to be a contradiction to the above mentioned detoxification process. This can, indeed, be the case if certain conditions are not considered carefully. Degussa has developed a new process to enable the application of H2O2, in this field without having any detrimental effect on the cyanide. The incentive to seek an oxygen supplier like H2O2 instead of air was the fact that one of the most important conditions for the fast dissolution of gold is the presence of physically dissolved oxygen in the pulp (Kydryk and Kellog, 1954; Warstell, 1987). 2 Au + 4CN- + 1 /2 O2 + H2O => 2 [Au(CN)2]- + 2 OH Complete gold recovery within a reasonable leaching time may only be achieved if the pulp is saturated with dissolved oxygen. Gold losses in leach operations mainly occur due to the following reasons: • In some cases, the saturation of oxygen in the pulp cannot be achieved by aeration due to the difficulty to transfer oxygen from the gaseous into the liquid phase. This happens, for example, in pulps with a high viscosity. In such cases, even violent aeration only results in a low concentration of dissolved oxygen in the pulp. • In any case, the starting phase of leaching requires the highest amount of oxygen. Therefore, it is especially important to raise the dissolved oxygen level to saturation in the starting phase. Peroxide assisted leach process (PAL) The intention of the process is to introduce oxygen in a liquid form. As a result of intensive research work, it could be shown that hydrogen peroxide can be used as a source of oxygen in the leaching of gold ores. To avoid side reactions, the following conditions must be applied: • Hydrogen peroxide must be diluted before it is added to the pulp. • The dosage of H2O2, must be properly regulated based on the oxygen level of the pulp. • An important part of the continuous 02-control is a special oxygen electrode, which resists the extremely abrasive action of the pulp.

Jan 1, 1989

By Russell L. Wood

Introduction It is the morning after and there is one monumental hangover among many, if not most, of the oil companies that invested in mining in the past decade. How an oil company looks at mining depends on the time period. A few years back, around the time of investment, they were probably looking at a balance sheet of their own that was beginning to go wild. Many of their commodity prices already increased two-, three-, or fourfold, with no end in sight. There was a need to invest their earnings someplace before the stockholders or the government got too interested. And that someplace needed to be an industry where their own management skills could be applied. In mining there was an industry amazingly similar to their own. Both worked with nonrenewable natural resources, dealt with the same types of government agencies, and had the same type of critics and pressures from the general population. Oil executives saw mining as an industry that was somewhat backward as compared with their own. They also saw how a generous ration of the capital and management skills, so handsomely bestowed upon the oil industry, could be used to great advantage. But, looking at mining through the eyes of an oil company today, this is what might be seen: • An industry that has been spectacularly unprofitable for the past few years; • An industry of many commodities, any of which can go bad on short notice for protracted periods, for a variety of unpredictable and uncontrollable reasons, or for no reason at all; • A labor-intensive industry, often beset by labor problems. And with only limited potential for further economic mechanization or automation in today's primitive mining environment; • An industry confronted with external, foreign, and political problems, with no reasonable expectation of near-term relief; and • An industry that does not appear to have many "up" periods in its immediate future. And, there is little chance for either big, fast bucks or long periods of large, steady earnings. What Went Wrong? The visions of a few years back seemed as correct then as today's viewpoints seem to be now. What happened to change these views? What went wrong? The simplest and easiest answer is that if the economy was normal back then, nothing would have gone wrong. This is an easy answer but, unfortunately, probably incorrect. Let us take a broad overview. Certain events enabled and prompted oil companies to invest heavily in mining. These same events were major factors in precipitating the worldwide recession that has grievously injured mining and has brought those mining investments into disrepute. Taking a somewhat narrower, but still broad overview, the simi-

Jan 3, 1985

By M. Gobla

"Since 1994 the Bureau of Reclamation has used risk analysis to evaluate dams. Risk considers both the likelihood of failure and the expected consequences should failure occur. Risk analysis relies on the detailed formulation and critical analysis of all possible potential failure modes. Potential failure modes are formulated by describing the sequence of events that must occur to result in failure, where failure is defined as the uncontrolled release of the impounded reservoir. Failure occurs in response to loadings such as water seepage through a dam (static forces), shaking from an earthquake (dynamic forces), sudden flood inflows (hydrologic forces), or accidental overfilling (operational errors). Reclamation has applied risk analysis methods to inactive and abandoned mines to evaluate tailings dams and flooded underground mines. Case histories of potential failure mode analysis as applied to mine impounded water illustrates what critical technical information is needed to determine if failure is likely, what the consequences of failure are likely to be, and what are the highest risks that should be addressed first. INTRODUCTION Impounded water, and high water content wastes, at mining and industrial processing sites are potential threats to public safety and the environment. Recent failures such as the TVA Kingston flyash impoundment failure in Tennessee in 2008 (5.4 million yd3 released), the Mt Poly Tailings Dam failure in British Columbia in 2014 (5.8 million yd3 released), and the Smarco Fundao Tailings Dam failure in Brazil in 2015 (78 million yd3 released) has renewed attention to this issue. Potential risks from impounded water at mining and industrial sites originates not only from tailing dams, but also from water in process ponds and conveyance pipelines, sediment retention dams, water retention dams, mine pit lakes, and flooded underground mine workings. Conventional water retaining dams also have a record of failure. Notable examples of failure include Vajont Dam in Italy in 1963 (2,600 lives lost), Banqiao and Shimantan Dams in China in 1975 (171,000 lives lost), and Machhu-II Dam in India (10,000 lives lost). After the 1976 Teton Dam failure in Idaho, the Bureau of Reclamation (BOR) established a Dam Safety Program. By 1984 the agency developed and began using risk analysis methods. Reclamation operates and maintains an inventory of hundreds of aging dams which were designed and constructed over a period of one hundred years. Many of these facilities were constructed using engineering design and construction methods which would not be considered good practice according to present day standards. Diligent observational monitoring, detailed site investigations, and the application of risk evaluation to the BOR dam inventory has identified many deficiencies resulting in significant demands on available funding for remediation efforts. The application of risk analysis has allowed the agency to rank sites by risk. By giving priority of remedial efforts to high risk sites, significant improvements to dam safety have been achieved. BOR has also applied risk analysis to evaluate and reclaim inactive and abandoned mining and industrial sites. The application of risk analysis methods in the mining industry is not new, the methodology was transferred to the mining industry more than 15 years ago as discussed in ICOLD Bulleting 121 (ICOLD, 2001), yet failures persist. This paper is meant to convey the lessons learned from BOR’s use of risk analyses in its work on mine sites."

Jan 1, 2017

By W. E. Benton

Diatomite or diatomaceous earth is a siliceous sedimentary rock principally made up of the skeletal remains of the diatom, a microscopic, single-cell aquatic plant related to algae. The unique structure, chemical stability and physical properties are the major attributes of diatomite not met by other forms of silica. The production of diatomite was down in 1982 vs. 1981, reflecting the worldwide recession. The principal markets for diatomite products continue to be basically filtration and fillers. The outlook is for small but steady growth into the 1990s.

Jan 1, 1983

By Mark L. Matthews

In 1978, Congress enacted Public Law 95-604, the "Uranium Mill Tailings Radiation Control Act of 1978." The Act authorizes the Department of Energy to stabilize and control the tailings from designated inactive uranium mill tailings sites in a safe and environmentally acceptable manner. DOE has established the Uranium Mill Tailings Remedial Actions (UMTRA) Project Office to implement, in cooperation with the affected States, Indian tribes and owners of the sites, remedial actions at the sites. The UMTRA Project Office has initiated efforts to perform remedial actions in accordance with EPA cleanup standards.

Jan 1, 1981

By S. J. Page

The personal dust monitor (PDM) is a combined tapered element oscillating microbalance (TEOM) respirable dust monitor and cap lamp, configured to have dimensions and weight similar to those of a lead-acid miner’s cap lamp battery. The Higgins-Dewell (HD) cyclone design operated at 2.2 L min-1 airflow has previously been reported to have low bias relative to the International Organization for Standardization (ISO) convention. Previous reports have described the PDM as meeting the NIOSH Accuracy Criterion. In order to avoid certain analysis anomalies encountered during these evaluations, a different method of establishing the PDM relative standard deviation was employed. Re-analysis of laboratory data for both the early prototype and precommercial PDMs indicates that the PDM still meets the NIOSH accuracy criterion, and that the precommercial PDMs are more accurate than previously reported.

Jan 1, 2011

By K. L. Johnson

Networks of monitoring wells were established prior to mining at four longwall mine sites in the Appalachians to monitor the effect of mining on water levels, water quality and well yield. Two of the sites were located in stream valleys and the other two sites were located on hilltops. The depth to the mined seam ranged from 500-850 feet with the deeper cover being associated with the hilltop sites. Maximum declines in groundwater levels during and after mining were an order of magnitude less in the stream valley aquifers than in the hilltop aquifers. However, water levels in hilltop wells located outside the area immediately above the panel were unaffected. Water quality changes in wells as a result of longwall mining were negligible at all four sites. Increases in aquifer transmissivity and well yield were much more pronounced in the stream valley aquifers than in the hilltop aquifers. The significantly larger recharge area for the stream valley aquifers than that for the hilltop aquifers may partially account for the differing responses of these two classes of aquifers to longwall mining. These observations appear to be consistent with the stress-relief fracture model for Appalachian Plateau groundwater hydrology proposed by Wyrick and Borchers in 1981.

Jan 1, 1986

By L. C. Lang

INTRODUCTION The introduction of 165-mm (61/z-in.) holes to underground mining operations has made possible the application of Canadian Industries Ltd's (CIL) vertical crater retreat (VCR) mining method, illustrated in the accompanying sketches. This unique and revolutionary new application of spherical charge technology (see the Appendix), when applied to primary stopes and pillar recovery, eliminates raise boring, slot cutting, and dilu¬tion of ore by backfill; greatly improves fragmentation; reduces labor and time requirements; eliminates uphole drilling and blasting; and minimizes or completely elimi¬nates damages by blasts to the walls and retreating backs of the stope or pillar. If vertical large diameter holes are drilled on a designed pattern from a cut over a stope or pillar to bottom in the back of the undercut, and spherical charges of explosives are placed within these holes at the calculated optimum distance from the back of the undercut and detonated, a vertical thickness of ore will be blasted downwards into the previously mined area. Repeating this loading and blasting procedure, min¬ing of the stope or pillar retreats in the form of horizontal slices in a vertical upwards direction until the top sill is blasted and the mining of a stope or pillar is completed. The VCR method is also applicable to drop raises and has the potential to replace all other raising methods under most circumstances. PILLAR RECOVERY Levack Mine Inco Metals Co., Ontario Div., provided the first opportunity for the method in pillar recovery. The Levack mine's high grade ore pillar No. 4800 on the 975-m (3200-ft) level was used for the production¬scale experiment (Figs. 1 to 3). The pillar was about 49 m (160 ft) long, 6 m (20 ft) wide, and 20 to 26 m (65 to 85 ft) high. The mined area on both sides of the pillar was backfilled with a 1:30 cement:sand mixture. The pillar was removed in two phases. In phase 1, the standard uphole method was used to blast down the 18-m (60-ft) long section of waste from the bottom of the ore into the undercut. From the pillar's top sill, 165-mm (61/a-in.) holes were drilled downward into the pillar, and by measuring the depth of the holes, the results of the uphole blast were determined and roof line 1 was established. The bottom of each hole was plugged, then filled with sand to place the center of gravity of each spherical charge (loaded from the top sill) at a predetermined optimum distance from the horizontal free surface. The charges were then detonated. After detonation, both draw drifts were full of extremely well-broken material. The depth of each hole was measured again, and the plot of these depths re¬sulted in roof line 2. The same blasting procedure was repeated and the resultant new back elevation was marked by roof line 3. The poor results of the initial uphole blast at one location (notice the peak in area 1) appeared to influ¬ence the subsequent new backs. A third blast success¬fully evened the back, and resulted in roof line 4. An unblasted slab averaging 6.3 m (20.9 ft) thick remained below the pillar's top elevation as the final sill. In all three spherical charge blasts fragmentation of the blasted material was extremely fine. The backfill was fully exposed on both sides of the now-blasted pillar. The backfill remained undamaged and the ore was not diluted by sand. The remnants of all the 165-mm (61/2 -in.) holes remained clean and undamaged, and the holes had well-defined bottoms that could be easily measured and plugged. Each blast took down a 3.9 to 4.2-m (13 to 14-ft) thick horizontal slab of ore. Productivity was three times greater than that of the previously practiced cut-and-fill method. Since this was the first such experiment, blasting the remaining 6-m (20-ft) thick final slab was the sub¬ject of some deliberation. If the described method was repeated, we could have ended up with a 1.8-m (6-ft) thick sill unsafe to work on. It was therefore decided to blast the whole sill using two spherical charges prop¬erly placed in each hole, but with the application of

Jan 1, 1982

By Carl R. Christiansen

Underground sites have great potential for industrial plants, for warehouses, and for other purposes where above-surface locations are now employed. Subsurface locations have generally been over- looked by industrial planners even though they may offer many features of superiority over the conventional sites. There are many opportunities for industries or warehouses to go underground. New space can be excavated expressly for this purpose or openings in existing mines may be converted to useful space. Another possibility is the pre-planned mine-industrial site, in which a mineral product is mined and marketed as a commercial venture while utilizing a mining plan which will permit rapid conversion of the openings created into active industrial areas while mining continues in the adjacent ground. An underground plant located in chambers created by excavating space in solid rock has many advantages, including low construction costs, low heating-cooling costs, freedom from storm damage or other surface weather conditions, absence of traffic noise and vibration, fireproof construction, low insurance costs, excellent working conditions, easily controlled temperature and humidity, almost unlimited bearing capacity of floors, low maintenance costs, no effect upon surface.

Jan 1, 1973

Hoses provide a variety of critical functions in nearly every mining operation. Care must be taken when selecting, installing, and maintaining hoses to insure maximum hose life. This article provides some tips on selecting, installing, and maintaining hoses, and gives some background on hose construction and safety. In the mining industry, hoses are used for everything from hydraulic systems that power drills and hammers, to suction systems that carry away dust and debris. Hoses are used on such a large scale that a detailed discussion of specific applications would require many volumes. Dayco Corp. has, however, provided these basic and brief guidelines on how to select and maintain hoses for mining applications. Hose Construction Industrial hoses are classified as much by construction as they are by application. So, it is a good idea to start any discussion of hoses with a description of their basic construction. There are four basic components to hoses: the tube, reinforcement, cover, and couplings. The tube is the inner core or lining, and is usually made of smooth rubber or plastic. Rubber or plastic composition is scientifically determined to be compatible with the materials passing through it, such as fluids, chemicals, gases, or substances. Other considerations in compounding rubber or plastic for the tube are temperature ranges and stresses. In practice, the required compound can be precisely chosen from various charts listing them according to application. There are, however, applications where the compounding of tube rubber or plastic must be customized. The reinforcement hose layer immediately surrounds the tube. As its name implies, this layer gives the hose strength and provides protection for the tube component. Reinforcement must be able to contain the inner forces-pressure and shock-transmitted through the tube, and such external forces as twisting, bending, and crushing. So it is usually made of high tensile strength cord, fabric, or wire that is wrapped, wound, or braided around the tube. The cover is the outermost layer, so it may encounter oils, acids, water, dust, abrasion, and other potentially damaging influences depending, of course, on the application. An appropriate cover material must be selected, usually a rubber compound. Couplings, even though they are a hose attachment, are as vital as any other component and should be considered part of the hose. Couplings are usually made of carbon steel or malleable iron; however, bronze or stainless steel might be used if corrosive chemicals are involved and brass or aluminum might be used for flammable fluids. A coupling is usually in two parts: an insert that fits into the hose bore and contains the threaded male, female, or other type of connection, and the ferrule or collar that fits over the hose and secures the coupling in place. There are, however, some one-part couplings that fit securely into the hose without a ferrule. These are usually limited to low pressure applications and are secured to the hose by clamps or metal bands. Some fittings can also be permanently "built-in" as the hose is made. Couplings can be classified as reusable or permanent. The reusables can be removed when the old hose expires and used again on a new hose. The permanent types are crimped or swaged on at the factory or in the field (using special equipment) and must be discarded with the hose. Although couplings are most commonly thought of as having threaded fittings, there is another type called "quick connect." These either have a crow-foot or jaw-type design that permits a fast, "snap" connection. This can

Jan 6, 1982

By Martin L. Smith

Sampling and subsequent geostatistical structural analysis is being carried out on the Bunker Hill Superfund site in northern Idaho's Silver Valley. This paper reports on the findings and methodology used in the initial sampling campaign which had the goal of determining an appropriate sampling grid design, spacing and protocol. Emphasis is placed on those characteristics, of mine wastes which are unique and require special consideration when applying geostatistical analysis. This study represents one of the first applications of geostatistical techniques to the problem of hazardous mining waste site characterization. Remediation of Superfund projects similar to the one described herein represents opportunity as well as liability for the mining industry.

Jan 1, 1992

By John A. Herbst, William T. Pate, A. Edward Oblad

Dynamic optimization is a cost-effective means to increase production or improve product quality in a plant. Several model-based approaches to dynamic optimization of mineral processing operations have been developed and implemented in plants. This paper reviews the phenomenological models that are the basis for dynamic optimization and the alternative methods of mathematical optimization. These methods range from theoretically complex and costly to implement to intuitively comprehensible and readily implemented.

Jan 1, 1990

By Jerry R. Johanson

Introduction Many solids processes that look feasible on paper or even in pilot studies become disasters in full scale. Changes in solids flow characteristics from pilot scale to full scale be- cause of larger solids contact pressures, thermal shocks, or unforeseen temperature excursions often cause full-scale solids flow problems not encountered in the laboratory or pilot studies. Fortunately, proper applications of solids flow principals and solids flow properties measurement at full scale conditions can anticipate the problems and thus provide an accurate basis for full-scale design or existing plant modifications. Solids flow properties The basic solids flow property definitions are given in Table 1 (Johanson, 1966, 1976, 1978, 1985). In this section, the author will discuss each property and its potential for creating full-scale pilot plants. Unconfined yield strength The unconfined yield strength of a bulk solid determines the tendency for cohesive hangups of the solids in the process and is effected by the time at rest, chemical reactions, temperature, liquid content, solids size consist, and phase changes of solids or fluids involved in the process. One of the common mistakes is to consider the solids condition only at the beginning and the end of the process. For example, consider the direct reduction of iron ore. The feed solids are free-flowing, hard lump ore or pellets. The end result is a hard, free-flowing reduced sponge iron lump or pellet. One might conclude that the process involves free-flowing lumps. This conclusion may seem reasonable based on laboratory studies oreven pilot plant studies where solids contact pressures are very low. Unfortunately, during the reduction reactions at high temperatures, the rock or pellets soften and, with solids contact pressures associated with the full-scale plant, may fuse together, thus providing a possible solids hangup in the moving bed reactor not evident in the low pressure pilot scale. Fig. 1 shows some typical unconfined yield strength measurements for pellets as a function of solids contact pressure at various times at rest. Note at low consolidation pressures, generally associated with laboratory or pilot plant studies, the pellets show no strength unless held at rest for an extended time. This will produce a very free-flowing condition in pilot plants that will not exist in the full size reactor.

Jan 1, 1990

By Shawn Boltz, Erik Westman, Derrick Chambers, Xingdong Zhaoa, Qiankun Zhua

Microseismic passive velocity tomography is becoming an effective technology for large-scale mining-induced stress monitoring in underground mines, especially in the field of early warning of mining-induced hazards. Aiming at the problems of missing low-magnitude events in traditional microseismic monitoring, as well as the potential to increase the resolution of the tomograms in both time and space, a microseismic passive velocity tomography analysis method based on the template matching method was proposed. First, the template matching method was used to pick microseismic events. Then, the probabilistic, non-linear, global-search earthquake location method was used to locate the events and calculate travel time and travel distance. Finally, the fast-marching method was used to carry out passive velocity tomography analysis to image the movement of the abutment stress associated with the longwall. The results show that the selected templates can identify more microseismic events than the traditional microseismic analysis, and the variation trend of events in each day is consistent with the data from the traditional analysis. Using microseismic event locations, travel times and travel distance can effectively complete microseismic passive velocity tomography analysis in each day during the study period and can effectively identify the forward and side abutment stress zones of that longwall panel. This study provides a reference for the identification of mine-induced high stress zones, which has implications for hazard identification and mitigation, with increased resolution of the tomograms in time and space.

Jun 25, 2024

By L. Cui

X. Huang (1997) recently suggested a poroelastic model for simulating the hydraulic fracturing breakdown pressure. His paper began with a discussion questioning Haimson and Fairhurst's (H&F) model. He claimed that the H&F model failed to lead to the Hubbert and Willis (H&W) model as[ a --> 0.] Huang also tried to explain why the H&F model could only work for special types of rock conditions. He pointed out that one possible reason could be that the Terzaghi's effective-stress concept had been adopted. The H&F model (Haimson and Fairhurst, 1969) was derived under the conditions that the borehole wall is fully penetrated (pp = pi) and a drained state is realized (steady pore-pressure field). Cui et al. (1997a) demonstrated that, under drained conditions, the total stresses in the penetrating poroelastic model (identical to the H&F model) degenerate into their counterparts in the elastic model (identical to the H&W model) as [a -- 0.] However, the effective-stress conditions are different for both of these models, because different pore-pressure conditions at the borehole wall were adopted. In the H&W model, p = po was assumed, i.e., the pore pressure field is not disturbed; but pp = pi was assumed in the H&F model. Assuming that Terzaghi's effective-stress controls tensile failure (that was the hypothesis adopted in both the H&F and the H&W models), only the following two special drained poroelastic cases may degenerate into the H&W model for very small a: • when the pore pressure at the borehole wall remains at the same level as the virgin pore pressure for a penetrating model and when the borehole wall is simply impermeable (i.e., the nonpenetrating model, Cui et al., 1997b). Therefore, simply setting a = 0 in the H&F model generally does not lead to the same problem described by the H&W model. On the other hand, when the porepressure boundary conditions do not correspond to the ones in the H&W model, a degeneration of the poroelastic model to the H&W model as [a – 0] should be questionable. The pore pressure at the borehole wall is generally dependent on the injection-fluid pressure and the penetrating conditions at the borehole wall, such as the existence of a filter-cake. For an impermeable wall, pp is independent of Pi, and it is basically an unknown [(how¬ever, pp -- po as t --oo).] For a fully permeable wall, pp is the same as Pi. Between these two extremes, pp should be a function of p; and the permeable condition of the borehole wall, which may be dependent of the leak-off coefficient cf. (the range of cf is from 0 to 1). Theoretically, for a rock with low permeability, a penetrating borehole wall is still possible. For saturated porous materials with very low values of a, poroelasticity shows that a pore pressure will be built up due to the stress concentration subjected to a nonhydrostatic in situ stress field (Cheng et al., 1993). This phenomenon is known as the Skempton effect. The variation of the pore pressure may be evaluated by [AP = 3 B(A rr + 06ee + A (Y,,)] (1) where B is the Skempton pore pressure coefficient. This pore pressure variation dissipates as time increases (it is totally gone as the drained state is approached). The rate of the dissipation mainly rely on the permeability of the formation. The dissipation is very slow for tight formations because their permeability is very low, and it is fast for rocks of high permeability. According to our analyses, the time period for this process could be from seconds (for sandstone's) to a couple of days (for stales). One possible reason that the H&C model did not agree well with the experimental results for rocks with low permeability might be that the time interval between the application of the loading and the fluid injection had not been long enough for the dissipation of the Skempton effect. The effective-stress law basically defines how much pore pressure contributes to the total stress. The difference between Terzaghi's effective stress and Biot's effective stress is that 100% of the pore pressure contributes to the total stress in Terzaghi's definition, while only a certain portion of the pore pressure ((ap) goes to the total stress in Biot's definition. Therefore, when the pore pressure at the boretole wall (pp) is determined according to Biot's effective-stress law, app in the total tangential stress is attributed to the pore pressure. In other words, Terzaghi's effective tangential stress is expressed

Jan 1, 1999

By Paul C. Miclea

With the growing urban population and the concern for traffic congestion and pollution (emissions control), public transportation is becoming more and more attractive to both city dwellers and managers. To gain access to the central area of the city, the sub- way remains the most viable alternative, despite its higher cost when compared to above ground or elevated transportation systems. There are relatively few regulations and criteria for subway ventilation, particularly when compared with mine ventilation. The main document that provides guidance and general recommendations for subway ventilation and environmental control is the Sub- way Environmental Design Handbook, published in 1976 (2nd edition) by the U.S. Department of Transportation, Office of Re- search and Development. Many of the subway systems in existence today have been designed and built with ventilation features adequate for normal train operation, but their design does not consider stringent criteria for such emergency conditions as a train fire in a tunnel. The National Fire Protection Association's Standard for Fixed Guideway Transit Systems known as NFPA 130 as well as the ASHRAE Handbook - HVAC Applications (1995) provide specific design and operation requirements for subway ventilation systems. Some of the existing, old subways are upgrading their ventilation systems to comply with the new regulations. The ability of a particular ventilation system design to provide adequate ventilation during normal and emergency conditions can be evaluated using computer modeling and simulation techniques. A train fire will cause a sudden change in the tunnel ventilation pattern by adding an unsteady and fast-growing source of heat. The hot air and gasses created by a fire will tend to flow uphill, possibly against the normal flow, producing a 'backlayering effect." To prevent this effect from happening, enough ventilation must be provided and the governing criterion to establish the required airflow is called 'Critical Velocity." Several software packages are available for special applications on tunnel and station ventilation as well as to model the spread of smoke and heat in case of a major tunnel fire, using Computational Fluid Dynamics. This paper presents the current issues in subway ventilation, as shared by the public transit community around the world, together with modern alternative design concepts vis-à-vis the more stringent emerging Fire-Life Safety criteria. Examples of recent studies and ventilation systems design are provided.

Jan 1, 1997

By A. J. Kroha, N. Wesis

Most copper mines produce both ore and low-grade "leach" rock or acid waters that contain recoverable copper. The sulfide ores pre¬dominate, and a portion that is too low grade for milling to produce concentrates for smelting, but has to be mined and trucked away anyhow, may be leached successfully with acid in dumps. Most of this leach material consists of sulfides and silicates or carbonates, and if the gangue is such that it consumes a high quantity of acid, this factor may rule out a leach operation. There are also valuable deposits that contain mostly acid-soluble copper, or occasional sulfide ores from which a sulfide concentrate can be roasted and acid-leached to produce a copper-bearing solution. Finally, there are milling ores in which the lesser part of the copper is acid-soluble and can be precipitated with iron or synthetic inorganic precipitants that produce metallic copper or copper sulfides that will float with the sulfides. Ordinarily, ores that contain copper associated with the sulfur ion, such as in the minerals chalcopyrite, chalcocite, bornite (and others), are milled to produce a 25-30% Cu concentrate for smelting, while a lesser amount of acid-soluble copper may be converted from solution to cement copper on iron scrap. A fast-growing percentage of such copper, however, is removed from solution with exchange resins or organic compounds in organic carriers such as kerosene (solvent extraction), then eluted with strong acid and subjected to electrolytic precipitation either in marketable form or as anodes that can be refined further. From the point of view of conventional copper smelting, copper flotation concentrates and cement copper are of chief interest in this chapter. Table I is a condensed open schedule for concentrates that generally run between 25 and 35% copper, and much less frequently as low as 12-15% or as high as 65-75% copper, the former being due to intimate relationship with pyrite (like the former United Verde Extension), and the latter representing such ores as the Bolivian Coro¬coro ore in which the copper is in the form of chalcocite in sandstone. These extremes are no longer common. When they occur, a special purchase schedule has to be negotiated. Included in Table 1, copper precipitates (cement copper) generally run from 70-85%a copper, and the same basic purchase schedule is used as with flotation concentrates. Sulfide Flotation Concentrates The sulfide copper concentrate produced in the mill as a flotation froth, with water then added for transportation of the heavy mineral particles from the flotation cells to thickeners, may run 60-80% water by weight, and the removal of water down to 25-50% by weight by means of thickeners, followed by further dewatering by continuous vacuum filters to 7-18% moisture by weight (depending on size of solids by screen analysis and also by content of clay) is a critical operation. Mill operators would like to produce a filter cake with 7-9% moisture content, but even with the help of steam on the filter this desirable condition is seldom realized when the concentrate is as fine as 80% -325 mesh. More commonly, the final concentrate is reground in pro- to produce best copper recovery and grade of concentrate (or molybdenite separation). In those cases, increasingly frequent, the filter product may not be a cake at all, but a mud that is hard to handle-even requiring a thermal dryer. Greater difficulty of form¬ing a manageable cake often comes from the copper-molybdenum separation by flotation, because the alkaline sulfides and hydrosulfides, or cyanide, or other similar reagents used for the separation, may leave the now relatively molybdenite-free copper concentrate even more difficult to filter. Handling a wet filter cake is difficult enough when its destination is only a short distance away-a matter of yards rather than miles. In those cases the filter cake may be thermally dried near the point of production, using rotary or multiple hearth, or fluidized-bed dryers. Alternatively, the concentrate may be pumped or carried in slurry form to the smelter and filtered there, or it may be spray-dried and compacted. For transportation to a smelter just a few miles to a few thousand miles away by ship or railroad other factors may be important, such as: in shipping by sea, avoidance of spontaneous combustion; in shipping by rail, losses by leakage if too wet or by wind and sun if too dry. It is the responsibility of the millman-usually the mill superinten¬dent-to make sure that his concentrates are in satisfactory condition when they leave the mill so that they meet these requirements: 1) They must have been accurately sampled and dry-weighed, the latter meaning that a moisture determination and gross weight must have been taken. 2) They must be dried sufficiently when necessary to prepare them for safe transportation. 3) They must arrive at the smelter with reasonable likelihood that they can be check-weighed and sampled fairly and equitably,

Jan 1, 1985

By Victor C. Storhok, Batric Pesic, Marlyn D. Ballain

No1ihwest Alloys produces magnesium from dolomite by reduction with ferrosilicon and aluminum. This process produces alkaline by-products in quantities that allow their consideration for use as commercial products. Two such products are dicalci­um silicate, termed as Di-Cal, and magnesium oxide, termed simply MgO. The alkaline properties of these materials offer acid neutralizing power, comparable to traditionally used lime. Thus, the materials warrant study for their use in various appli­cations where neutralization of acidic waters or process solutions are required. One such application is the treatment of acid mine waters with the objective of removing heavy metals. The chem­ical stabilities of the produced metal precipitates are critically important, from an environmental point of view, for their safe disposal in tailing ponds. The EPA regulates waste material dis­posal according to the results from TCLP testing. In this ongoing study, three mine waters were selected for treatment with Di-Cal and MgO. Lime was used as the control for these experiments. The waters were collected from the Bunker Hill mine, Berkeley Pit, and the Blackbird Mine. In addi­tion to testing the chemical stability of produced metal precipi­tates by the TCLP procedure, two other parameters were studied as well. They are: 1. the effect of water pH on stability, and 2. the effect of pH on the stability of resulting precipitates. In the first three months of study, it was found that neutral­ization with Di-Cal results in precipitates that range from equiv­alent to superior in performance as compared to those produced using lime for the waters from all three mines. An additional beneficial characteristic of precipitation with Di-Cal is the fast filtering characteristics of produced precipitates, when compared to the precipitates produced by lime. This property may be of significance when considering the improvement in drainage of tailing ponds and the volume of solids produced. This paper will discuss in detail the stability of metal precipitates with respect to various pe1iinent metals found in mine waters of Berkeley Pit (Cu, Zn, Mn, etc.), Bunker Hill (Zn, Mn, Ni and Co), and Blackbird Mine (Cu, Co).

Jan 1, 1998