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By B. C. Paul, R. Q. Honaker, M. Huang

Froth flotation is now, the most commonly used method for cleaning fine-coal fractions in coal-preparation plants. Past research and industrial practice have demonstrated that froth flotation achieves excellent separation efficiencies for fine coals having only a small fraction of middling particles. However, the efficiency of froth flotation decreases sharply as the concentration of middling particles increases. This is due to the flotation of high-ash particles that have a small amount of hydrophobic coal present at the surface. Because gravity-based processes are much more efficient for the treatment of middlings, the application of centrifugal washing for fine-coal cleaning was studied with an emphasis on the primary cleaning of fine-circuit feed. Results are shown of a parametric study that used a semibatch Falcon centrifugal concentrator. The results of the study revealed that separation performance is a function of particle size, bowl geometry, feed solids content and feed rate.

Jan 1, 1996

By S. S. Hsieh

As part of a program to address the long-term need to ensure adequate supplies of domestic phosphate raw materials, the Tennessee Valley Authority (TVA) has conducted considerable research on dolomitic phosphate rock beneficiation. The discovery that alkyl diphosphonic acids act as phosphate mineral depressants led to the development of several processes involving a carbonate flotation step for the effective beneficiation of these problem ores. Most recently, a partial deoiling process in which the reagent consumption is greatly reduced has been developed. The reduction is achieved because part of the initial feed is beneficiated without passing through the separate carbonate flotation step. Using this method, a phosphate product containing 30.6-30.9% P2O5 and 0.7-0.8% MgO was obtained from a Florida dolomitic phosphate pebble that contained about 23% P2O5 and 2.0% MgO. The P205 recovery was 80.8-86.5%.

Jan 1, 1991

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

Finely disseminated iron ores compose much of today's reserves. After removing the silicate gangue, the resulting fine iron oxide concentrates require pelletization with a binder to enable transportation and reduction. Binders can be used more efficiently if their binding mechanisms can be completely understood. Typically, bentonite clay is needed as a binder to improve the strength of the pellets. Bentonite is an effective binder at a relatively low cost, but costs could be reduced further by replacing bentonite with even lower-cost binders, such as high-carbon Fly-Ash Based Binder (FBB). In this study, it was found that bentonite binds by a physical mechanism, while FBB binds by a chemical mechanism. These mechanisms were determined to be incompatible. Partial replacement of either binder with the other reduced dry magnetite pellet compressive strengths below the industrially acceptable value of 22.2 newtons (5 lbf). Activators and accelerators that activated the fly-ash deactivated the bentonite. The incompatibility of these two binders is explained.

Jan 1, 2000

By Stephen B. Castor

The rare earth elements (REE) which include the 15 lanthanide elements (Z = 57 through 71) and yttrium (Z = 39) are so called because the elements were originally isolated in the late 18th and early 19th centuries as oxides from rare minerals. Most REE are not as uncommon in nature as the name implies. Cerium, the most abundant REE (Table 1), comprises more of the earth's crust than copper or lead. Many REE are more common than tin and molyb¬denum, and all but promethium are more common than silver or mercury (Taylor, 1964). Promethium (Z = 61) is best known as an artificial element, but has been reported in very minute quantities in natural materials. Lanthanide elements with low atomic numbers are generally more abundant in the earth's crust than those with high atomic numbers. In addition, lanthanide elements with even atomic numbers are two to seven times more abundant than adjacent lan¬thanides (Table 1) with odd atomic numbers. The lanthanide elements traditionally have been divided into two groups: the light rare earths (LREE), lanthanum through eu¬ropium (Z = 57 through 63); and the heavy rare earths (HREE), gadolinium through lutetium (Z = 64 through 71). Although yttrium is the lightest REE, it is usually grouped with the HREE to which it is chemically and physically similar. The REE are lithophile elements (elements enriched in the earth's crust) that invariably occur together naturally because all are trivalent (except for Ce+4 and Eu+2 in some environments) and have similar ionic radii. Increase in atomic number in the lanthanide group is accompanied by addition of electrons to an inner level rather than the outer shell. Consequently, there is no change in valence with change in atomic number, and the lanthanide elements all fall into the same cell of the periodic table. The chemical and physical differences that do exist within the REE group are caused by small differences in ionic radius, and generally result in segre¬gation of REE into deposits enriched in either light lanthanides or heavy lanthanides plus yttrium. The relative abundance of individual lanthanide elements has been found useful in the modelling of rock-forming processes. Comparisons are generally made using a logarithmic plot of lanthanide abundances normalized to abundances in chondritic (stony) meteorites. The use of this method eliminates the abundance vari¬ation between lanthanides of odd and even atomic number, and allows determination of the extent of fractionation between the lanthanides because such fractionation is not considered to have taken place during chondrite formation. The method is also useful because chondrites are thought to be compositionally similar to the original earth's mantle. Europium anomalies (positive or negative departures of europium from chondrite-normalized plots) have been found to be particularly effective for petrogenetic modelling. REE were originally produced in minor amounts from small deposits in granite pegmatite, the geologic environment in which they were discovered. During the second half of the 19th century and the first half of the 20th century, REE came mainly from placer deposits. With the exception of the most abundant lanthanide el¬ements (cerium, lanthanum, and neodymium), individual REE were not commercially available until the 1940s. Since 1965, most of the world's REE have come from two hard rock deposits: Mountain Pass, United States, and Bayan Obo, China. GEOGRAPHIC DISTRIBUTION OF REE DEPOSITS More than 70% of the world's REE raw materials come from three countries: China, the United States, and Australia. China emerged as a major producer of REE raw materials during the 1980s, while Australian and United States market share decreased dramatically (Fig. 1). Table 2 gives recent annual production figures along with estimated reserves by country, and Fig. 2 shows loca¬tions of significant REE mining. MINERALS THAT CONTAIN REE Although REE comprise significant amounts of many minerals, almost all production has come from less than ten minerals. Table 3 lists minerals that have yielded REE commercially or have po¬tential for production in the future. Extraction from a potentially economic REE resource is strongly dependant on its REE miner¬alogy. Minerals that are easily broken down, such as bastnasite, are more desirable than those that are difficult to dissociate, such as allanite. In general, producing deposits contain REE-bearing min¬erals that are relatively easy to concentrate because of coarse grain size or other attributes. For more thorough discussions of REE¬bearing minerals see Mariano (1989a) and Cesbron (1989).

Jan 1, 1994

By S. Atak, G. Bulut

The xanthate adsorption mechanism on pyrite was investigated with an emphasis on solubility studies. Additionally, adsorption studies, Eh and infrared spectrophotometer measurements were performed. Pyrite floats at pH levels between 3 and 6 using ethyl and amyl xanthates with a maximum at pH 4. The solubility of pyrite in the absence and presence of xanthate was investigated in detail with respect to pH and conditioning time. Experimental results showed that the solubility of pyrite increases from 60 to 120 ppm (10-3 to 2x10-3 mol/ L) in the presence of xanthate at pH 3. Fourier transform infrared (FTIR) spectra of ferric xanthate, dixanthogen and pyrite conditioned with xanthate indicated that dixanthogen is the only species that can be determined on the pyrite surface. Adsorption tests showed that the highest level of xanthate adsorption was in the acidic range; yet higher adsorption densities were detected in neutral solutions compared to alkaline conditions. Electrochemical data suggest that the doubling of the iron concentration is due to the reduction of ferric ion to ferrous by xanthate while xanthate is oxidized to dixanthogen. Dixanthogen formation together with the hydrophobic iron hydroxy-xanthate complex on the surface is proposed to be responsible for pyrite flotation. Key words: Flotation, Pyrite, Adsorption, Xanthate, Dixanthogen

Jan 1, 2003

In the hierarchy of performing mine planning tasks, production scheduling is naturally at the bottom. Yet good production scheduling is probably the most demanding and difficult task to achieve for various reasons. Being the last task in the mine planning hierarchy, a proper production schedule must not only conform to both the long-range and the short- range mine plans but also satisfy many practical details that are unique to day-to-day operations. For example, the averages which were good enough for both long-range and short- range mine planning are no longer good enough for production scheduling. Instead, the variations become the overriding concern. The answer must also be obtained quickly and on time, thus adding another dimension to the overall problem. Because of these considerations, many production scheduling problems have been begging for solution during the past decade. Many attempts were made to solve the production scheduling problems in the past using computers and the available operations research (OR) techniques, particularly among academic groups. However, it is probably true that, compared with the current full-time computer usage in long-range and short-range mine planning, very little computer usage is being made in production scheduling to date.

Jan 1, 1979

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'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 Richard L. Profio

A practical analysis of the three basic elements of a mining shovel's productive capacity. Payload Speed Reliability Improved dipper fill factor is explored referring to the latest generation of wide dippers combined with higher "bank horsepower" capabilities of modern drive systems. Cycle time improvement is reviewed including overall attachment geometry and coordination of motions. A specific case study is detailed. Basic reliability of welded structures, machinery components, and electrical systems are related to operating costs. Three basic elements determine a mining shovel's long term productive capacity: 1) Average payload per cycle. 2) Average cycle time; and 3) Reliability. This paper deals with productivity gains that are possible with existing shovels as well as new equipment.

Jan 1, 1984

By Anthony G. Fonseca

About 45% to 50% of the coal mined in the United States passes through coal preparation plants. East of the Mississippi River, this number increases to about 75% to 80%. Although the cost for coal preparation is worthwhile to some, the coal industry faces the challenges of continuing downward pressure on the price of coal and the impact of new environmental regulations. Coal preparation is an effective process that achieves 75% to 80% ash reduction, 15% to 80% trace element reduction and 85% to 90% Btu recovery. It is less effective for pyrite reduction (35% to 70%) and on-line operating time (40% to 60%). It also suffers from obsolete control systems. In coal preparation, raw coal is modified to a more preferred form to enable its use. The methods used to achieve this modification must be cost-effective and appropriate to the raw coal, the scale of production and the end-user requirements.

Jan 1, 1995

By F. L. Pierson

Since the discovery of 'potash in the Carlsbad area in 1926 and the start of mining operations in 1931, "first mining," operations have been carried on. This is a room and pillar system in which approximately 62% of the total ore is removed. Or the basis of ore reserves and rates of extraction, the life of the mine in the early; days was estimated to be well over 50 years, however, even with an increase by additional ore reserves, it became increasingly apparent that the life of the mine was being shortened considerably by increased production. Further cause for the shortening of mine life was due to the exceptionally weak back over the northeast quadrant of the main orebody which had .caved in on first mining up to a height of 8 feet. Thus, in 1953 the decision was' made to begin pillar pulling, also called, second mining or final mining, and in December of 1953 this program was started after much discussion.

Jan 1, 1965

By Z. Kowalski

In the non-operational ?Górka? quarry, there are situated both a waste disposal site of an area of 6.7 ha, containing approximately 1 million ton wastes from high aluminium content materials (incl. about 400000 t of wastes from aluminum hydroxide production), and a pond of an area of roughly 3 ha, up to 15 m deep, with about 400000 m3 of polluted water and infiltrates from the disposal site mentioned. The basic problem related to the effluents in the Górka pond is both their high alkalinity (pH 12-13.4) and a variable analysis of pollutants whose level increases appreciably with the pond?s depth. The solution offered provides, first of all, remedial pumping and a treatment of effluents retained in the Górka reservoir. Effluents would be treated in a reverse osmosis plant according to a process verified so far on a quarter-commercial scale. The treatment process by-product, a brine solution (containing about 25% NaCl) whose amount may reach up to 100000 m3 might be supplied to the Debiensko Mine, which produces technological salt from salt-bearing mine waters. In parallel, an intake of underflows of infiltrates would be provided (previous geochemical identification) in order to take them away from the high aluminum site (by means of deep wells, and possibly girdling ditches). Such effluents would be subject to treatment. The successive stage is the utilization of approx. 50000 m3 of bottom slurry, left in the emptied reservoir, in the production of self-solidifying mixtures, which would be then used for sealing the bottom and the banks of the Górka reservoir. The empty and sealed reservoir would be filled up with a volume of approx. 600000 t of ashes, coming from coal burning in the thermal-electric power station. For the last stage, instead, is scheduled the rehabilitation of the entire waste disposal site in the Górka quarry. The site will be drained, covered with a leveling, sealing and rehabilitating layer, which will allow to introduce plants on the rehabilitated waste site. Later on, the area could be turned into a town park

Jan 1, 2006

By Andrew T. Zador

The growing demand for energy combined with the need to protect and preserve the environment have become prime considerations in the design of efficient coal handling ports and their associated stockpiles. The transportation system is analogous to a connected chain which originates at the mines and terminates at steel mills or power plants, in which the port and the stockpile reserves are a vital link. Railroad transportation costs can be minimized by rapid discharge of the cars which, in turn, requires that an adequate stockpile system be available. Ocean transport costs can be minimized by having an efficient high capacity ship loading system and sufficient quantities of coal on hand -all of which also requires an efficient stockpile system to meet shipping demands. The connector between the two modes of transportation, therefore, is the port stockyard. It is here at the interface that optimization of land utilization and materials handling equipment results in a total transportation chain that can operate as a smooth efficient system. Multiple origins of coal require the simultaneous handling of large quantities of coal of many grades. Railroad transportation logistics often limit and define the incoming arrival patterns and limited rolling stock often require priority treatment for railroad operations. The physical limitations of prime ocean frontage require that coal stockpiles be built close to residential or industrial areas which entail strict environmental controls. Large capacity, enclosed coal stockpiles for ocean shipping have not yet been considered feasible but may be forthcoming in the near future. Several have been analyzed.

Jan 1, 1978

By Werner Wenzel

Whisker growth during the reduction of iron ores The phenomenon of whisker growth can be observed in a number of solid-state crystallisation reactions, which, apparently, appear in6 the first instance as though they are two independent processes. From such reactions, one can find the whisker type of crystallisation on solid bodies. These examples show, that the whisker growth is neither due to certain crystal forms with predetermined temperature and pressureranges, nor to parameters, which, ad hoc, are necessary for the crystallisation. In all cases the whisker formation seems to depend on nucleation and the availability of crystallizing material in direct environment of such a nucleus, under given parameters of a system. One can generally state, that the research of whisker growth with the aim of establishing proper physical laws, to determine the foundamentals of whisker formation and whisker growth, is at present under active study. We can conclude that, although a great number of hypothesis have been proposed; no concluding theory has yet been arrived at. In the institute for ferrous metallurgy of the technical university of Aachen, research works are undertaken to find an explination to various metallurgical phenomena through help of whisker growth which was observed under varying conditions of iron ore reduction. This was possible by using refined microscopic techniques through which the whisker growth on solid oxyde substrates is detected. This refined technique combines the latest methods of surface examination with the help of stereo-scan-electron-microscope and proper and carefull preparation of surface to be stereo-scanned.

Jan 1, 1971

By Alfred Petrick, Roger Dewey

INTRODUCTION Estimation of capital and operating costs for mines and mills has long been a necessary ingredient of feasibility studies, and though many practitioners have successfully used "rules of thumb" and factor estimating methods for many years, a great deal of work is usually required to prepare an initial estimate and/or to update or revise such an estimate. Development of comprehensive mining and milling cost models during the 1970's by the U.S. Bureau of Mines and by Canadian workers (O'Hara and Mullar) provided the basis for feasibility-level cost estimates, but the numerous equations and the apparent rigor and complexity involved in preparing a project estimate using these models has discouraged many, if not most, potential users. Their acceptance was also hindered by a broad misunderstanding by many engineers in the minerals industries who had little exposure to such procedures, and by a possibly inadequate level of explanation or documentation of the parametric approach to cost estimating. The preliminary nature of most databases, geologic/engineering/ and market, available to the engineer or analyst during the early stages of a mineral project is a universal problem. Basing corporate mega-investment decisions or even individual equipment selection decisions on the usual "SWAG’s has always been a risky practice, but in the current business environment, the "guesstimates" are almost 'verboten', and are certainly becoming antiquated with the advent of both the described models, the ubiquitous microcomputer, and the nearly-ubiquitous hybrid species known as the "computerized engineer". Engineers have almost always sought techniques of cost estimation that can be applied and justified when geologic and engineering information for a mining project is preliminary and incomplete. The answer, in some cases, is the "parametric" approach. Equations, based on actual, past experience are developed to relate capital and operating costs to major physical variables such as the rate of mining or milling, with modifiers for such details as mining method, process characteristics, methods of construction, or types of equipment . Cost Engineering Standards Five major types of fixed capital cost estimates have been defined by the American Association of Cost Engineers : 1. Order-of-Magnitude estimate based on previous cost data and minimal knowledge. usually a "ratio estimate" with no better than +/- 30% accuracy. 2. Factored or ratio estimate based on major equipment costs with probable accuracy in the 25 to 30% range. 3. Budget authorization or Preliminary estimate, based on sufficient project data to

Jan 1, 1987

By Roy Moreby

Dartbrook Coal is currently developing a modern, high production (3.5 Mtpy) single longwall retreat coal mine in the Upper Hunter Valley of NSW, Australia. The location is a green field site, with no surface or underground mining operations ever having taken place in the initial target seams. The coal seams of the working section, floor and immediate roof average a total 20 m in thickness and contain a mixed carbon dioxide 1 methane seam gas ranging from 6.5 to 11.0 m3/t. The target seams have undergone detailed characterisation to assess hazards and quantify the controls required to minimise associated risks during mining. Risks identified during initial phases of exploration and 14 months of development include: • High gas emission during development. • Probable high longwall gas emission rate. • Medium to high propensity to spontaneous combustion • Possible carbon dioxide outbursts on geological structures • The main controls currently in place include, • Mine geometry and ventilation circuits • Exploratory drilling ahead of development • Intensive in seam pre drainage • Goaf drainage plants • Comprehensive monitoring systems • Underground environmental management plan. • Pre mining predictions of gas emission provided an indication of gas emission rates which have proven to be an under estimate. Current mine issues are; • Both surface fans are at full capacity (250 m3/s 1.4 kPa) • Both gas drainage pumps are at full capacity (2.3 m3/s) • Alternatives to extensive in seam pre drainage being investigated (surface to in seam hydro fracture) Review of mine design underway This paper describes the seam characterisation, assessment of mining hazards and summary of the management plan developed to address the risks identified. A review of current mine underground environmental performance is also provided

Jan 1, 1997

By William W. Wellborn

Abrasives are substances both natural and synthetic that are used to grind, polish, abrade, scour, clean, or otherwise remove solid material usually by rubbing action (as in a grinding wheel), but also by impact (pressure blasting). Facts about the discovery, early use, adaptation, and development of abrasives prior to the 20th century are few. From the sparse, earliest information that is available, it would appear that the science of abrasives began when, to fashion his tools, man began to select certain kinds of rocks in preference to others because of their peculiar properties. There is evidence that stones were sawed by some sort of grinding machine as early as 4000 BC in Egypt. The real expansion of abrasives use coincided with the beginning of metallurgy in the Middle East around 2000 BC. The earliest confirmed use of abrasives for grinding metal comes from a steel dagger found along with a sharpening stone which was dated at about 1500 BC. Natural and manufactured abrasive substances play an important role in fashioning and finishing numerous products with a wide variety of end-use applications. The mining and preparation of crude, natural abrasives for their markets and the manufacture of synthetic and other abrasive materials contribute significantly to the US economy (Table 1). In addition, just as sawing to specific dimensions enhanced the value of the Egyptian stone, the use of abrasives in fabrication of modem products, and the addition of abrasives to others, add value. GENERAL CONSIDERATIONS The most important physical properties of materials that qualify as abrasives are hardness, toughness (or rigidity), grain shape and size, character of fracture (or cleavage), and purity (or uniformity). To fabricate bonded abrasive products such as grinding wheels, additional considerations include stability under high heat and bonding characteristics of grain surfaces. The economic factors of cost and availability are always important. No one singular property is paramount for any use. For some applications extreme hardness and toughness are needed, as in diamonds for drill bits. For other uses, factors of greatest importance are hardness and ability to break down slowly and, in the process, develop fresh cutting edges when grains become worn. In garnet sandpaper, for example, highly cleavable, friable, or extremely tough grains are not desirable. For other uses, extreme hardness may be objectionable. Abrasives for dentifrices and for glass-cleaning soaps are cases in point. For the most efficient uses in more critical applications, different types of abrasives are rarely completely interchangeable. While crushed quartz and garnet are both used in sandpaper, the two abrasives are not at all interchangeable in their applications. In the final analysis, the choice of a high-grade abrasive depends upon the quality and quantity of work performed by the abrasive per unit of cost. Initial cost of an artificial abrasive may be much greater than that of a natural abrasive, but the artificial mineral may do so much better work and do it so much faster that the ultimate cost is less. It is for this reason that artificial abrasives have largely replaced natural abrasives.

Jan 1, 1994

By H. Rush Spedden

A process has been developed and piloted to reject copper, lead, iron, phosphorus, bismuth, antimony, elemental sulfur, and insol from molybdenite concentrate without decomposing the contained molybdenite and rhenium. The impure concentrate is first heated with concentrated sulfuric acid at 260 to 290°C for 1 hour. This bake not only oxidizes the contaminating sulfides, but also activates the concentrate for later rejection of insol by flotation. Two stages of leaching follow the sulfation step. The first is an acidified water leach which extracts the oxidized copper, iron, and phosphorous; the second leach rejects the lead, bismuth, and antimony into a warm concentrated solution of sodium chloride. Final upgrading is achieved by conventional froth flotation. The chemistry of the process as well as results obtained from its implementation in a 5,000 lb/day pilot plant are presented.

Jan 1, 1971

By B. K. Parekh, R. H. Mehta

Handling of fine coal is an importance issue for coal as well as the utility industry. Reconstitution in form of pellet or briquet would be desirable if it could be done economically. This paper evaluates the effectiveness of three binders e.g., asphalt- emulsion, corn starch and Brewex, in forming pellets of ultra-fine clean coal. It was found that asphalt emulsion and corn starch were not effective binders for ultra-fine clean coal, however, Brewex provided excellent quality of pellets, which exceeded all the minimum quasi requirements of coal pellets. INTRODUCTION Advanced coal preparation techniques (physical or chemical) are designed to process ultra-fine (minus 200 mesh or 74 pm) coal. The average particle size of the clean coal product is about 25 pm. Currently available dewatering techniques are effective in providing a 25 to 30 percent moisture product of this size coal slurry (Parekh et al. 1989). Handling and utilization of this dewatered ultra-fine size coal is difficult. Reconstitution, which includes pelletization, briquetting, extrusion, etc., of the dewatered coal product provides an ideal answer to handling and transportation problems. However, reconstitution of fines in the coal industry has not been successful due to various reasons. The most important criterion for its unsuccessful application in the coal industry is the cost of the binder (Capes et al. 1965). Various methods of reconstitution, such as briquetting, disk pelletization, extrusion pelletization and roller and die-pelletization, have been tested for fine coal agglomeration (Conkle et al., 1991) (Richards, 1990). Conkle et al., (1992) have provided a detail analysis of each technique in terms of moisture and operating cost requirements. According to Conkle, briquetting requires a product containing very low moisture, whereas pelletization can be done with coal containing up to 30 percent moisture. Currently available dewatering technique provides about 25 percent moisture product, thus, pelletization appears to be an ideal route for reconstitution. In this paper, results are presented on the comparison of three different binders i.e., asphalt-emulsion, corn starch and Brewers condensed solids (BCS or Brewex), for fine coal pelletization.

Jan 1, 1995

By Leonardo Rosas Sánchez, Darío Ángel Bulla

"The two track road sector between Cisneros-Loboguerrero (having a length of 14 km and 11 short unidirectional tunnels) is a new highway still on construction located along the existing Highway Buga – Buenaventura (Colombia). Due to the steep topography, difficult geological/geotechnical conditions, the traditional portal construction approach was impossible to be implemented in eleven (11) of the twenty two (22) portals. In this specific project, the skewed portal approach was developed as the optimal and more effective solution. The construction time and costs of portals were reduced in 21 months and 75%, respectively. At the same time, the natural equilibrium of the slopes was maintained, thus, facilitating adequate safety factors along the useful life of the project. This document pretends to show a detailed design experience (and the construction of the entry portal of tunnel 11 (left)). INTRODUCTION The Cisneros-Loboguerrero project is a 14 km length dual lane highway, connecting productive zones with the Pacific Ocean along the current Bogota-Buenaventura highway as shown in Figure 1. The project is located at the northern side of the Dagua River Valley. The preliminary design suggested constructing portals under challenging topographical conditions, i.e., slope cuts having heights between 50 to 100 m (some having slopes > 100%) requiring complex developments and having very high risks.The entry portal of tunnel 11 (left) was the first portal successfully built by implementing the skewed portal approach. Later, this approach was used as a model for the construction of the other ten (10) portal having similar topographic and geotechnical conditions. The development of this new alternative provided a considerable reduction on the construction time of portals, thus relieving a critical path in this project."

Jan 1, 2016

By D. Malhotra

The number of variables which can effect flotation of molybdenite is very large. This paper discusses the effect of two operating variables, namely, the degree of aeration and agitation, on the grade and recovery of molybdenite. A series of factorially designed flotation tests were carried out in a "Leeds Automatic Laboratory Flotation Cell." The analysis of flotation results indicate that molybdenite can be characterized into fast-floating and slow-floating components. The rates of fast-floating components, which do not create a problem in flotation, could be varied by changing agitation of airflow rates. However, the rate of slow-floating components was stable and was only a function of residence time in the float machine. The degree of aeration has a significant effect on the grade of MoS2 with only a slight effect on the recovery of MoS2. The degree of agitation has a slight effect on the grade, but does not effect the recovery of molybdenite. The implications of the results on the design of flotation cells and control and optimization of the existing circuits is subsequently discussed.

Jan 1, 1977