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By H. Walqui, A. Buendía

The Toquepala concentrator expansion project incremented the capacity of the existing concentrator by 15,000 Metric tons per day (t/d), from 45,000 to 60,000 t/d. The plant is located in southern Peru,1,200 km southeast of Lima. The implementation plan for the expansion project was based upon an engineering and procurement contract, and a construction contract. The overall Project Management was by Southern Peru Copper Corporation (SPCC). The expansion followed a very aggressive schedule that considered a construction period of only ten months, with minimum interference to the existing facilities. Major equipment was purchased during basic engineering. This paper details the design, construction, and commissioning of the 6.4 m diameter by 10.2 m long ball mill, and the performance of the milling circuit during the first two months of operation.

Jan 1, 2003

By Joseph F. Shirley

Introduction Inorganic depressants are chemicals added to the froth flotation process to maintain or produce hydrophylic surfaces on one or more minerals so they will remain in the pulp rather than attach to the air bubbles. Whether a depressant will be effective in depressing a given mineral depends on how it is used and the environment in which it is used. Some depressants must be added and conditioned into the pulp before the collector is added. Other depressants can be added in the presence of previously added collectors and effectively depress the desired minerals. Often the concentration of depressant relative to the collector concentration is critical to making an economic separation. It should be noted that some depressants will only work when specific collectors have been used. Some reagents such as sulfuric acid and hydrogen peroxide essentially act as depressants by altering the previously added collectors so that they are no longer effective. Some depressants such as sodium ferrocyanide are only effective in a very limited pH range and an essentially collector free environment. Major Depressants CaO Milk of Lime is by far the most used depressant both in tonnage and dollar value. Although it is most commonly listed as a pH modifier, it is also an important depressant for pyrite in copper and copper molybdenite flotation plants. Although a simple pH adjustment will often depress pyrite, there are times when grinding with fresh milk of lime is required to obtain optimum depression. Lime is also a molybdenite and gold depressant. NaSH Sodium hydrosulfide is by far the most commonly used reagent to depress copper and iron sulfide minerals in by-product molybdenite flotation plants. NaSH also depresses gold, silver sulfides and silver chloride. ZnSO4 Zinc sulfate is commonly used to depress sphalerite in lead-zinc and copper-lead- zinc flotation plants. It has also been used to effectively depress talc in by-product molybdenite circuits at a pH of 6.5.

Jan 1, 1986

The next 13 years gave me an education that anyone, then or now, would well envy. W.T. Reid had founded Belmont School along the lines of Groton and, as at Groton, he sent many graduates to Harvard. He obtained excellent teachers and maintained as strict a discipline as the independent nature of the western students would stand. With a beautiful location in an unspoiled little valley in the foot- hills of the Coast Range there was every chance for walks and expeditions. The years at Belmont were good ones. My closest friends were Harold Jordan and Lewis Hackett. Lewis was son of the military director at the school. With half a dozen other very young inmates we early won the sobriquet of "The Dirty Dozen." Sliding down a steep muddy hill on barrel staves helped us earn the name. An inherited shyness exaggerated my sensitivity at being a "teacher's boy" and this prevented a close association with the wilder western boys who made up most of the schoolmates. Trapping wildcats and coyotes in the brush-covered hills and taking long walks gave me plenty to do. I had grown too fast and when I went to college was 6 ft 3 in. tall and weighed only 145 lb-hardly a respectable bean pole. I was only two months past 16 when I left for Harvard. The long summer vacations were happy ones. Most of them were spent in the high mountains, and even when I was ten years old I could keep up with anyone on long steep climbs. I learned to handle a rowboat during our stays at Lakeside, a rustic resort at the south- east corner of Lake Tahoe that could then be reached only by the steamer that made a dally trip around the Lake. Several of the teachers at Belmont remained good friends as long as we were on the same side of the continent. The finest was C.H. Rieber, who taught mathematics at Belmont for the first few years we

Jan 1, 1976

By Robert S. Hedin, David M. Hyman

Acid mine drainage (AMD) is a common water pollution problem on active and abandoned coal mine sites in the eastern United States. AMD forms when surface mining brings unweathered pyrite-containing rocks to the surface or when deep mining allows oxygen to contact buried pyritic strata. In the absence of neutralizing compounds, the drainage that results can be extremely acidic and contaminated with dissolved iron, manganese, and sulfate. Drainages with pH < 3.0 and concentrations of sulfate greater than 1,000 mg/L, iron greater than 50 mg/L, and manganese greater than 10 mg/L are common. Where water flows through alkaline materials (such as limestone) before surfacing, the drainage is less acidic and occasionally circumneutral, but it can still contain high concentrations of sulfate and metals. Current water quality standards in the United States require that mine discharges have a pH between 6 and 9, total iron concentration less than 3.0 mg/L, and manganese less than 2.0 mg/L. At thousands of active and inactive mine sites, drainage does not meet these standards and is being treated before discharge by the mining company. At thousands of other sites, which were abandoned prior to the enactment of water pollution laws or were operated by companies that have gone bankrupt, untreated AMD is polluting receiving water systems. The standard mine drainage treatment system involves the addition of a1 kal ine chemicals to the water, which raises the pH and cause metals to precipitate in a settling pond. These systems are expensive, often costing tens or hundreds of thousands of dollars per year for chemicals, operation, maintenance, and disposal of the metal- laden sludge. Because the drainage on many sites will likely be contaminated for decades, there is financial incentive to find alternative water treatment systems. In the last five years, many mining companies and engineering firms have experimented with the construction of wetland treatment systems. Very few of the several hundred wetland systems that have been built are performing well enough to justify total abandonment of chemical treatment. Nonetheless, systems continue to be built because the mining companies have found that wetlands can reduce their AMD treatment costs. It is also their hope that further research and development will eventually result in systems that will completely rep1 ace chemical treatment and offer long term performance at minimal cost. The Bureau of Mines has been actively involved, from a research perspective, with this approach to AMD treatment since its inception. In this paper, the status of constructed wetland technology is discussed with respect to the construction and performance of systems, chemical and biological processes that affect AMD chemistry within constructed wetlands, and the future of this technology as perceived by the Bureau of Mines. ORIGINAL OBSERVATIONS The constructed wetland concept has its roots in observations of natural Sphaqnum peat wetlands that received acid mine drainage and, instead of being adversely affected, appeared to cause an amelioration of the polluted water (Huntsman et al. 1978; Wieder and Lang 1982). These observations instigated the idea that wet1 and systems might be used for the intentional treatment of mine drainage. Because the discharge of AMD into a natural wetland is prohibited in the U.S. by several laws, it has been necessary to construct wetlands which act solely as water treatment systems. Initially, most wetland research and construction efforts mimicked the original observations by utilizing Sphaqnum moss and peat. Despite promising laboratory results (Kleinmann et al. 1983; Tarleton et al. 1984; Burris et al. 1984; Gerber et al. 1985), virtually all field tests of Sphasnum-dominated constructed wetlands failed to provide sufficient water treatment for more than several months. Sphasnum proved quite sensitive to the stresses associated with transplanting, abrupt changes in water chemistry, excessive or insufficient water depth, and

Jan 1, 1989

By A. J. Lynch, M. J. Lees

Introduction The design of wet grinding circuits may vary considerably according to the duty which is to be performed. Grinding machines which are commonly used in these circuits are rod mills, ball mills, autoge¬nous mills, and pebble mills while size separation machines commonly used are vibrating and wedge-wire screens, hydrocyclones, and rake classifiers. In some circuits, such as those producing feed to cement kilns, all the ore entering the circuit is discharged in a single product stream. In others, such as those operating on magnetite-taconite ores, the feed entering the circuit leaves in two or more streams. Grinding circuits in flotation plants normally include sections in which either the concentrate or the tailings from the primary flotation circuit is reground and these additional sections should be regarded as an inte¬gral part of the grinding circuit. A typical circuit is shown in Fig. 18. Problems in achieving maximum efficiency may occur at three stages: 1) At the design stage when the optimum size reduction flow¬sheet must be selected with respect to the types, numbers, and sizes of processing units. 2) At the operating stage when the correct selection of values must be made for those variables which may be altered while the circuit is off-line but which are constant while the circuit is on-line, for instance the vortex finder diameter of the cyclones. 3) Under conditions of continuous operation when changes in the circuit feed or in the mechanical operation of the circuit may cause undesirable changes in the circuit product. The first two prob¬lems are concerned with optimization, that is, with making the best choice from many possibilities, and the last with control. Optimization Problems. The main purpose of a wet grinding circuit is to produce particles with a sizing distribution which will ensure optimum performance of the later chemical reaction or physical concentration circuits, because it is on the efficiency of these later circuits that the economics of the total mineral treatment process depend. The size distribution of the product from a grinding circuit may be affected by many variables, some concerned with the arrange¬ment and operation of the machines and some concerned with the ore characteristics, and it is important that these should be recognized in the design and operation of the circuit. For instance, in a simple ball mill-cyclone circuit the variables which affect the product size even for constant feed characteristics are shown in Fig. 19. In addition, the rate, size, hardness, and mineralogical composition of the feed will affect the performance of the circuit and the product size. When the circuit contains more than one mill, both the number of choices available and the complexity of the problem are increased consider¬ably. Consequently, one type of problem which is to be solved if an optimum product is to be obtained from an existing grinding circuit is that of circuit design, and this includes arrangement of processing units and selection of those operating conditions which cannot be altered when the circuit is operating, such as mill speed and cyclone inlet and outlet dimensions. The other major problem in optimization is encountered at the design stage when a choice must be made about the type, number, and size of processing units. For instance, for a typical high-capacity size reduction operation, should two circuits each containing one rod mill and two ball mills or three circuits each containing one rod mill and one ball mill be used? Alternatively, should ball mills only be used and if so, how many, what size, and what arrangement?

Jan 1, 1985

By J. A. M. Cano

With a view toward increasing its product competitiveness in the foreign market, SAMARCO MINERACAO S/A has developed high MgO type pellets at its pilot plant for performing direct reduction simulation tests. Additives such as dolomitic lime and magnesite were used for increasing pellet MgO content. Comparative direct reduction simulation tests were conducted both in SAMARCO and foreign laboratories. The test results showed that it is possible to produce direct reduction type pellets with satisfactory metallurgical properties, which can be successfully processed at increased temperatures in direct reduction furnaces.

Jan 1, 1992

By S. K. Singh

"Feather edging"-a major source of injury and fatal accidents in Australia, South Africa and USA, refers to a phenomenon where the roof, at the goaf edge, falls as a thin wafer of rock often suddenly, over-riding conventional timber breakerlines. It may run down and affect adversely a working bord at distances of 15 metres or more from the goaf edge. Feather edging typically occurs in relatively massive, strong roof strata such as sandstone or conglomerate. This paper presents the results of a mine trial where a concentrated row of fully grouted bolts were able to resist the feather edging in the first comprehensive, monitored trial in an Australian pillar (&apos;total&apos;) extraction coal mine.

Jan 1, 1999

By J. R. Gump

The term "rare earths" is used to designate a group of elements, Numbers 57-71 which are closely related chemically. They were called "earths" because their oxides resemble those of the alkaline earths, and the designation "rare" was originally applied because of their scarcity. Now many of these elements are known to exist in considerable quantities, and the term "Lanthanide" is applied to this group of elements, because of their similarities to lanthanum, the first element in the group. Many authorities insist on including yttrium, element Number 39, because of its extremely close association with the members of the group in nature and because of its similarity in chemical properties. It has been estimated that the combined rare earth group makes up about five-thousandth. of one percent of the Earth&apos;s crust, a higher percent than the common elements, lead, tin, zinc, arsenic, mercury, silver and gold. Cerium is by far the most abundant member comprising about half of the total amount. Yttrium, neodymium, and lanthanum are next in order; samarium, gadolinium, praseodymium, dysprosium, erbium, holmium, and ytterbium are increasingly scarce; while the elements lutetium, thulium, terbium, and europium are extremely rare. The rare earths as a group are metals, usually trivalent, forming basic oxides, and insoluble oxalates in dilute mineral acids. Their fluorides are also difficultly soluble; hence the rare earths or lanthanides may be separated, in general, from other elements by adding oxalic or hydrofluoric acid to a dilute mineral acid solution containing them. The most striking fact which characterizes these elements is the remarkable .similarity in both the physical and chemical properties of their compounds. Their main differences are in the solubilities of their salts and the basicity of their oxides, which vary between that of the alkaline earths and that of aluminum.

Jan 1, 1960

By D. Hodouin

The improvement of mineral processing data using material balance computer programs is nowadays a current practice in many plants, engineering companies or research centers. It now becomes more obvious that such programs should integrate new features such as the preprocessing of data, the estimation of less accessible process variables and an in-depth interpretation of: the computed results. A few solutions to these new problems are already available. The feasibility of a mass balance computation for instance can be examined by analyzing the degree of redundancy of the measurements prior to any data adjustment. New variables such as chemical analyses in particle size intervals or liquid phase composition can be included in a set of measurements. Finally, a sensitivity analysis can be used to analyze the reliability of the results, to detect eventual outliers or to redesign a sampling scheme. Some illustrations of these new developments are presented.

Jan 1, 1989

By Mogens Henriksen

In the past decade, numerous urge scale drilling projects have made use of reverse circulation air lift methods. Reverse air lift pumping can be established by injecting gas into a column of liquid; the subsequently decreased density of the gas-liquid mixture allows atmospheric pressure outside the lift conduit to establish flow. The rate of flow can be controlled by varying the gas to liquid ratio. A typical reverse circulation air lift system is depicted in Figure 1. Allen1 has described a number of reverse air lift drilling projects. Although cited examples are generally related to large diameter holes, the method has seen use in other applications such as sand dredging, chemical pumping, water well pumping and continuous coring. One particular benefit obtained with enclosed reverse circulation drilling is elimination of lost circulation. The required air/water ratio for pumping at a given submergence and lift has been established experimentally by Gibbs2. His data, however, do not reflect sensitivity of flowrate to air/water ratio. Further, in Gibbs&apos; work, air apparently is injected continuously along the column of water. This configuration differs from that used in reverse air lift drilling where air is injected at discrete points only. Gas lift theory has received considerable attention over the past decades. Orldszewski3 defined flow regimes in two phase flow. Griffith and Wallis4 determined relative velocities of gas to liquid. Two-phase flow has been studied extensively for oil transport in inclined and vertical pipes by Chierici et a15, Beggs and Brill6, and Gould, Tek and Katz.7

Jan 1, 1977

By H. J. H. Pieterse

This paper highlights the mineralogy of the Barrick Goldstrike Mines Inc. area. The mine is located at the north end of the Tuscarora Spur and is part of the Bluestar/Goldstrike subdistrict of the Carlin Trend. The Betze, Post, and Meikle Mine are major deposits located in the area. The deportment of gold, its leachability, and various process options are also evaluated. Particular emphasis is given to the practical and economic considerations of ore control, and routing of ore to a process option.

Jan 1, 1993

By Mark W. Cadwallader

Heap leach mining requires liners in order to contain leaching solutions for both economic and environmental incentives. The use of HDPE geomembrane liners for heap leaching is discussed in this paper. Comparison is made to clay liners, with inclusion of economic considerations. Several heap leach projects incorporating geomembranes are discussed.

Jan 1, 1987

By B. S. Crocker

Introduction Pebble mill grinding can be defined as wet or dry grinding in a tumbling mill with pebbles used as grinding media instead of metal balls. To clarify the terminology, which may differ in the various mineral industries, it is essential to point out the differences between grinding with partially sized and carefully sized grinding media. When grinding with partially sized media (limited top size only), sizing is usually the setting of the primary crusher and there is no bottom sizing-all fines, chips, sand, and small sizes are included. This is often called run-of-mine or primary autogenous grinding. In this case the entire crusher discharge is fed into the mill and part of it becomes "charge" and the rest is "sand feed." When grinding with carefully sized media (i.e., media that is sized both at the top and bottom of the size range), no fines or chips should be present in the charge to the mill. This is pebble milling and the grinding media consists of ore, rock, Danish pebbles, gravel pebbles, or ceramic pebbles-media that is appreciably lighter than steel balls. The grind¬ing charge is fed to the mill to maintain a load and it does the grinding usually on a classified sand product of suitable size. If the grinding media is the same as the raw material being ground, this type of grinding is also called secondary autogenous grinding. Pebble milling was in common use in the gold mines of South Africa in the early 1900s. It is by far the oldest form of autogenous grinding. It may be used to grind crushed ores from a maximum size of about -% in. to all -10 µm. Pebble milling differs from "run of mine" grinding in many ways, but two big differences are: I) The ore itself is crushed to approximately t/4 in., or it may be further reduced to 8 or 10 mesh in a rod mill or fine crusher. 2) Then carefully screened ore of the correct size is placed in one or more stages of pebble mills to grind the ore to its desired final size. The media size is carefully selected to be the most efficient for the particular size reduction required in each stage. This is a precise practice, and pebble mills can be designed as easily and accu¬rately as can rod mills or ball mills by those skilled in the art. A pebble mill is a rotating cylinder in which the grinding media has been carefully sized for the job in hand. If the mill is doing coarse grinding, the media will be correspondingly large and the mill might be called a primary (or intermediate) pebble mill or a lump mill. If it is to do fine grinding, the media selected is proportionately smaller and the mill would be called a secondary or tertiary pebble mill. The media itself is normally rounded pebbles made from screened ore from the primary-crusher discharge. The charge could be rounded pebbles of the correct size removed from a primary autogenous mill, or the pebbles could be carefully sized gravel, sized beach pebbles, or sized flint pebbles. Usually the specific gravity of a pebble is from 35 to 55% of that of steel balls. It commonly varies from 2.66 to 4.2, depending on the specific gravity of the ore. Primary pebble mills, or lump mills, would be charged with pieces of rock up to 10 in. maximum and 5 in. minimum and would grind crusher discharge with a maximum size of 1/4 in. on hard ores and 3/8 in. on softer ores. Secondary pebble mills could be charged with pieces of rock screened on 4 x 4-in. square mesh and retained on 1 1/2 x 1 1/2 in. These mills would grind material with a top size of 8 to 10 mesh. Tertiary pebble mills could be charged with screened rock that would pass 3 x 3 in. and be retained on 1 x 1 in. and grind 35 to 48 mesh sands. Principles of Operation Media Size To operate a pebble mill efficiently it is imperative that great care be taken in screening the rock feed that is fed into the mill to round up into a range of pebble sizes that are suitable for grinding of the sands that are also being fed into the mill, having regard to the top size and the average size of such sands. Since most mill operators know from experience the correct size of steel balls to be used in a mill to handle different sizes of sand feed, it is convenient to relate the pebble size to a ball size. If the correct ball size is not known, it is relatively simple to conduct batch laboratory tests in 12- to 18-in. laboratory mills to determine the optimum size of steel ball. Also, Bond and others have published formulas for calculating the optimum ball size. When the correct size of the individual steel ball is known, then a pebble of the same weight will perform the same grinding if used in the proper size of mill and under the proper conditions, which will be defined later. This media size, as defined by the average weight of the individual pieces, is most important. The following tests will serve to clarify this point. Laboratory Tests Using Steel Balls or Pebbles. Several labora¬tory grinds were made to compare a given size and weight of steel ball with a pebble of the same weight but larger size, and of the same size but lighter weight. The following loads were used in the 12-in. laboratory mill: All 1 1/4-in. steel balls weighing 122 ± 10 g. All 1 3/4-in. pebbles weighing 123 ± 5 g. All 1 1/4-in. pebbles weighing 45 ± 3 g. To equalize capacity, the test with the steel balls was run only 34% as long as the tests with pebbles. In each case, 2,500 g of primary sands were ground at the optimum gravity for the media being tested at a speed 78 1/2 % of critical speed, and for 170, 340, and 680 revolu¬tions for the steel balls, and 500, 1,000, and 2,000 revolutions for the pebbles. The first test is summarized in Table 74 by showing the 80% passing size in microns for the feed and for each of the grinds. Two other tests are summarized in Table 75 showing the screen analyses

Jan 1, 1985

By Ronald E. Feltman

Most engineering firms engaged in the design and manufacture of milling components and systems have established, through the years, a relatively well defined system design procedure. The process design engineer has at his disposal several alternative approaches to the successful completion of this task. Unfortunately, while these alternatives, at first glance, appear to be rather straight Forward, any attempt to implement them into a fully integrated commuter design program becomes increasingly more complex when one addresses himself to the problem in total. This paper summarizes Kennedy Van Saun&apos;s approach to the solution of this problem through a planned system design. Following the problem definition and discussion, a detailed analysis and system design synthesis is presented; the major aim here being to provide a system design which may be programmed and run on a relatively inexpensive computer installation such as the IBM 1130 system currently in use at Kennedy Van Saun. A summary section is also included, which Provides a general outline guide to the program systems design of equipment which, while quite different from milling system functionally, nevertheless lends itself to the approach used in this paper.

Jan 1, 1969

By W. M. Ma, Daniel W. H. Su, K. Centofanti, Yi Luo, W. L. Zhong, Syd S. Peng

9.1 INTRODUCTION There are approximately 70,000 abandoned coal mines nation¬wide, which is about 35 times the number of underground coal mines presently operating. The US Bureau of Mines estimates that there are over 8 million acres of undermined land due to the extraction of coal, metals, and nonmetals. Subsidence has affected more than 2 million acres, and more than 99% of the subsidence is related to underground coal mining. There is reason to believe that some of the remaining 6 million acres of the undermined land have a high potential to subside. The expansion of housing, highways, commercial structures, and other facilities has required the use of many areas that are underlain by abandoned coal mines, and this growth will continue. Many subsidence problems are derived from the collapse of aban¬doned underground coal mines. There are numerous abandoned mine workings in the anthracite fields of northeastern Pennsylva¬nia, in the Appalachian bituminous fields, the Illinois Basin, the Rock Springs, Wyoming area, and other areas of the United States (Gray et al., 1976). Various room and pillar patterns of mining have been used in the dipping anthracite seams and the nearly flat-lying bituminous seams with considerable variation in the per¬centage of coal extracted. The progressive deterioration of pillars, mine floors, and mine roofs after long exposure to air and water may later result in the collapse of strata over the mine entries, the crushing of the remaining coal pillars, or the bearing failure of the mine floor beneath the coal pillars. Subsidence then results as the collapse reaches the ground surface in the form of differential strains, depressions, cracking of the ground, and sinkhole devel¬opment. Subsidence over active longwall mines, which occurs concurrently with mining or is completed within a short period following coal extraction, has been studied extensively over the past decade. On the other hand, subsidence over abandoned coal mines receives little attention by the researchers, mainly because it is difficult to predict and takes place decades after mining has ceased. The techniques of investigating the subsidence events over abandoned coal mines are similar to those employed for active mines except that at the outset, it is necessary to determine whether or not the subsidence events are mining-related (Chugh et al., 1986; Cummings and Singh, 1986; Peng and Hsiung, 1986). This calls for the identification and confirmation of abandoned mine workings under or near the affected surface structures. Gen¬erally old mine maps, if available, are acquired, the surface bore¬holes are drilled for confirmation of the accuracy of the mine maps and determination of the potential for continued subsidence. Sub¬sidence monuments are established and periodic surveys con¬ducted to determine the amount and trends of surface movement in and around the affected surface structures; surface boreholes are used to investigate the integrity of the underground structures (i.e., roof, coal pillars, and floor) by TV camera. They are also used for monitoring the vertical and horizontal movements of the subsur¬face strata by Sondex (FPBX) and inclinometer (PFBI), respec¬tively, for determining the continuity of the subsidence events. Tape extensometers, crackmeters, etc., are used to monitor the development trends of major cracks in the structures or on the ground. These data are used to identify the causes of the subsidence events. Finally, abatement methods are selected to stabilize the structures. It must be noted that most subsidence events over abandoned coal mines are reported after the fact, and investiga¬tions are begun some time after reporting, that the subsequent measured movements are generally much smaller than that at¬tained in the active mines, and that due to lack of knowledge about the damage conditions and the exact location of the abandoned mine workings with respect to the affected surface structures, the precise causes of surface subsidence or surface structural damages are in most cases very difficult to identify. 9.2 TYPES OF SURFACE SUBSIDENCE According to Gray et al. (1977), after examining 354 incidents of subsidence above abandoned mines in the Pittsburgh metropol¬itan area, the subsidence features have a mean diameter (i.e. the average of long and short dimensions) from less than 1 ft to 1600 ft, with 84% less than or equal to 15 ft; the subsidence features have a depth ranging from less than 1 ft to 48 ft, with 89% less than 25 ft; 66% of the subsidence features are deeper than they are broad. Nearly 59% of the subsidence features occur with overbur¬den less than 50 ft thick and 81% less than 100 ft thick. No subsidence features occur with overburden thicker than 450 ft (Fig. 9.1). Occurrence of subsidence incidents varies from imme¬diately to more than 100 years after mining (Fig. 9.2). In analyzing the characteristics of approximately 3000 chim¬ney subsidence features along the Colorado Front Range, Matheson and Eckert-Clift (1986) examined historical aerial photographs on 4- to 14-year intervals between 1937 and 1967 and found that the majority of observable surface subsidence features occurred within 30 to 40 years after mining. According to Gray et al. (1977), the most prevalent subsi¬dence features over abandoned mined land are sinkholes, with depths of more than 3 ft, and troughs or sags, usually less than 3 ft deep. Sinkholes are steep-sided pits, while troughs are shallow depressions much wider in area than sinkholes. A. SINKHOLE SUBSIDENCE A sinkhole is caused by the collapse of a mine roof that works its way upward. If it is not arrested during the process it will eventually reach the surface and emerge as a sinkhole. The process is governed by the thickness and character of the overburden, the width and height of the mine openings. Sinkholes are usually 3 to 20 ft deep and may be 2 to 40 ft in diameter, although most are fewer than 16 ft across (Gray et al., 1977; DuMontelle and Bauer, 1983). Newly formed sinkholes have steep sides with straight or bell-shaped walls. At times, they appear to be conical in profile with the apex upward. If the topsoil collapses, the top portion will widen to form an hour-glass shape. Sinkhole subsidence usually occurs over abandoned mines less than 165 ft deep (Hunt, 1979). Matheson and Eckert-Clift (1986) found that chimney sink¬holes are likely to occur when the ratio of overburden thickness to mining height (h/m) is less than 4 to 5. When h/m is between 5 and 10 to 11, the potential occurrence of chimney sinkholes decreases rapidly. When h/m is more than 10 to 11, less than 10% of the mine openings that collapse will induce sinkholes on the surface.

Jan 1, 1992

By X. K. Jiang

Upgrading molybdenite ore can be technically difficult and expensive if talc is a major impurity in the ore. The layered structure of molybdenite and talc decreases the surface specificity of floatation and suppression agents, thereby making molybdenite ore beneficiation inefficient if under some wet processing methods. However, oxidic talc surfaces may be distinct from the sulfidic or oxysulfidic molybdenite surfaces, such distinction related to electron work functions and/or densities of electron donors/acceptors sites. If such differences are present, the upgrading of finely sized molybdenite ore by dry triboelectrostatic methods may hold advantages over traditional wet methods.

Jan 1, 1999

By M. Kyzeridi

The paper presents a study on the reclamation of two mined-out bauxite open pits in the area of Kefali, Itea, Greece. The mining area has been reclaimed partially in the past, but its proximity to the town of Itea and a national road with tourist traffic, rose the interest to examine various scenarios for improving the landscape and the aesthetics of the site and introducing alternative land-use for recreational activities. Geospatial data were modelled using a Geographical Information System (GIS) that made it possible to study alternative proposals in terms of final landscape profiles, volumes required to be moved and the associated costs.

Jan 1, 2002

By Eugene M. Trisko

Presentation Overview State-Of-The-Art Scrub/Switch Analyses > Allegheny Power System > PSI Energy UMWA Phase I Coal Interests > 16 Phase I Scrub Plant Coal Profiles > Coal Price/Emission Allowance Interaction Southern Company Case Study > 2+ Million Ton Phase II Allowance Purchases > 10+ Million Ton Phase I Coal Switch > Georgia PSC Proposed Fuel Price Cap > Shift/Avoid Georgia Scub Costs Through System Allowance Purchases

Jan 1, 1993

By J. Hanna, H. R. Gonzalez

Four Ca-bentonitic clays from Alabama were evaluated for use in diammonium phosphate (DAP) or urea ammonium nitrate (UAN) fertilizer suspensions. The viscosity and gel strength of suspensions prepared with different proportions of a Western bentonite clay and sodium-exchanged Alabama clays were measured as a function of clay content and dispersant concentration. The results obtained suggested that three of the Alabama Ca-bentonites would replace about 40% of the more expensive Western bentonite in UAN and DAP fertilizer suspensions. The performance of the Nu-modified clays, prepared according to the Tennessee Valley Authority soda ash method, was as good or even better than those observed for the Mg-, NaIMg- or K-modified clays. The TVA clay modification procedure appears to be reliable and cost effective.

Jan 1, 1991

By G. A. Gunderson

Since our time is very limited, I would like to relate to you as best as I can a few important areas that I think may be helpful. 1.) The Real Difference between Dense Phase and Dilute Phase Pneumatic Conveying: Within the many variations of so called Dense Phase conveying, it is often difficult to determine to what degree a system is Dense Phase or if, in fact, it is a Dense Phase system at all. A system that appears and is generally known to be Dense Phase can under certain operating conditions perform as a Dilute Phase system. If, for instance, a high pressure system, which is generally a Dense Phase system, operates at 60 PSIG, say 4 atmospheres, utilizes enough air volume, the conveying velocity will easily exceed the normal design velocity of a typical Dilute Phase system. Now, on the other hand, a low pressure system which we will term a Dilute Phase system, operating at 15 PSIG; say 1 atmosphere, can convey a given product at well below the saltation velocity, like a Dense Phase system.

Jan 1, 1977