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By Luis Mercado

El presente texto analiza el panorama global nacional del carbón en el Perú, así también señala las posibilidades de este como alternativa energética frente al petróleo, debido al alza de precios de producción y venta. Para ello, se revisa las reservas de carbón, el contexto geográfico y geológico de las mismas y los métodos de explotación de los países industrialmente desarrollados.

May 1, 1986

By L. J. Buter, D. J. Collins

Introduction FMC's Paradise Peak gold mine was discovered by staff geologists in July 1983. Located south of Gabbs, in western Nevada, the deposit contains about 11 Mt (12 million st) of ore grading at 3.6 g/t (0.105 oz per st) of gold and 129 g/t (3.75 oz per st) silver. It is estimated that the ore contains 34 t (1.1 million oz) of recoverable gold and 933 t (30 million oz) of recoverable silver. The 3.6 kt/d (4000 stpd) mill began operation April 14, 1986, just 18 months after Davy McKee completed a feasibility study and construction, and two-and-a-months ahead of schedule. It also came in about 10% under budget at a cost of about $100 million. Crushing Run-of-mine ore is delivered from the mine in 45-t (50-st) haul trucks to a 272-kt (300,000-st) capacity open air stockpile. Ore is reclaimed using 5 m3 (7 cu yd) front-end loaders that feed a 91-t (100-st) hopper equipped with a 610- x 610-mm (24- x 24-in.) grizzly. A 1.4- x 11-m (54-in. x 35-ft) apron feeder delivers ore from the hopper by way of a vibrating grizzly screen with 127 mm (5 in.) openings to a 1- x 1.3-m (43- x 51-in.) Rexnord jaw crusher. Crushed material is conveyed to a 1.8- x 4-m (6- x 14-ft) Link Belt vibrating screen equipped with a 50-mm (2-in.) manganese grizzly bar deck. Screen oversize is fed to a 1.7/0.6-m (5.5 ft) Rexnord standard cone crusher. Product from this crusher joins the -50 mm (-2 in.) screen undersize and is conveyed to a 136-t (150-st) surge bin. Three hydraulically driven belt feeders deliver ore from the ore bin to three 2 x 6 m (7 x 20 ft) Link Belt single-deck tertiary screens fitted with 6 x 25 mm (0.25 x 1 in.) slotted polyethylene screens. Screen undersize - -6 mm (-0.25 in.) rock - is conveyed to a 3.6-kt (4000-st) ore storage bin. Screen oversize is delivered by way of a conveyor to a 91-t (100-st) surge bin. From this bin, two belt feeders feed ore to two 1.7 m (66 in.) Rexnord Gyradisc crushers, which are in closed circuit with the tertiary screens. Design features include magnetic protection for tramp iron removal and sloping floors and pump sumps for ease of clean up. Dust scrubbers keep the working environment clean and in compliance with state dust emission regulations. Grinding Two 1 x 4.5 m (3.5 x 15 ft) belt feeders deliver ore from the fine ore bin to a conveyor belt that feeds it to the 5 x 7 m (16 x 22 ft) Allis-Chalmers ball mill. The conveyor belt is equipped with a Merrick weightometer. This controls feed rate to the ball mill by speed control on the hydraulic motors driving the feeders. The 2.6 MW (3500 hp) powered mill grinds 3.6 kt/a (4000 stpd) to an average grind of 85% -75 µm (-200 mesh). The mill is in closed circuit with four 660 mm (26 in.) Krebs cyclones. Three of these are normally in operation while one is a standby. Warman pumps feed the cyclones through a cyclone distribution manifold. The cyclone underflow returns to the ball mill, which has a 300% average circulating load. Mill discharge pH is 11, and that is obtained through pebble lime addition to the feed belt. Cyclone overflow is sampled using a Heath & Sherwood two-stage sampler. It then flows to a 27-m-diam (90-ft-diam) Eimco thickener. Thickener overflow is returned to the grinding circuit, while underflow at 50% solids is pumped to the leaching circuit. Grinding-circuit design features include a steeply sloped floor for ease of clean up, an open area over the pumps for maintenance accessibility, and a ball hopper with a self discharge bottom to speed ball addition. Other features include a mill liner handler and an air clutch that connects the mill motor to the mill.

Jan 12, 1986

By G. H. Lewis, C. L. Krcil, N. S. Harding, D. E. Havekotte

Introduction Limestone injection work was completed in a 440-kW (1.5-million Btu per hour) pilot-scale combustor in preparation for a 50,000-kg/h (110,000-lb per hour) industrial steam boiler limestone injection test program. The main objectives were to: (1) identify the important variables that affect sulfur capture, (2) determine the maximum SO2 removal, and (3) determine the effect of limestone addition on furnace operability and heat transfer. A more detailed report is given by Lewis et al. (1985). The test program was divided into two phases. The objectives of Phase I were to determine the important variables that affect SO2 reduction and to investigate the effect of limestone addition on combustor operation and heat transfer surfaces. The objective of Phase II was to determine, using only the important variables identified during Phase I, the conditions that resulted in the greatest sulfur dioxide reduction. Phase I Test program Using one limestone, five variables were identified as having an effect on SO2 reduction. These are noted with their respective ranges: heat release rate [103 to 166 kW/m3 (10,000 to 16,000 Btu per hour-cu ft)], Ca/S molar ratio (0 to 3), flue gas oxygen concentration (3% to 5%), additive particle size [90% -150 µm (-100 mesh) and 90% -45 µm (-325 mesh)], and burner stoichiometry (SR = 0.75 and >1.15). A 2.3% sulfur northern West Virginia coal, pulverized to a nominal grind of 70% -75 µm (-200 mesh), was burned for the tests. The limestone used was obtained from a Maysville, KY, quarry and contained 51.75% (wt) CaO. An 11-run statistically designed program (plus three duplicates) was used to screen the five variables. The effect of limestone injection on the combustor heat transfer surfaces was determined by comparing the heat transfer rates through a slag panel and fouling probes during periods with and without limestone injection. Combustor operation was determined by noting any operating problems that occurred during the tests. Test procedure After furnace wall temperatures had stabilized at operating conditions, heat transfer surfaces (slag panel and fouling probes) and the convection section were blown clean, and a four-hour test was begun. The

Jan 12, 1986

By D. M. Avant, A. J. Solomon, S. M. Pickering, W. R. Fox

Introduction Most high purity silica is produced from crystalline quartz, generally from sand. However, there is a small area of southern Illinois where a porous, friable white chert rock is found that is very different in physical form from quartz sand. This microcrystalline material, commonly referred to as tripoli, is more easily ground to micron sizes than sand. Although of similar Moh's hardness, microcrystalline tripoli is less abrasive than either ground quartz sand or many synthetic abrasives. Location The southwestern corner of Illinois is underlain by extensive deposits of very porous chert; in some places the material is quite white. Two companies, Illinois Minerals Co. and Tammsco Inc., mine and process this white material. The Illinois Minerals plant is located at Elco, and the Tammsco plant is in Tamms, IL. Both companies mine from various locations within a 10-mile truck haul of their plants. Production from this district has been continuous since at least 1915. Malvern Minerals Co. mines and processes a generally similar microcrystalline, friable novaculite near Hot Springs, AR. Geology The southern Illinois white tripoli area is on the southern flank of the Illinois Basin. The commercial silica zone is a facies of the Clear Creek chert of Lower Devonian Age. To the north, where it occurs deep in the basin, the Clear Creek is a siliceous limestone. In the mining area, thin overburden has allowed access by surface ground water, which has leached away the carbonate content and left a quite pure silica residue. Fossils characteristic of sea bottom deposition, such as brachiopods and trilobites, are occasionally encountered in mining. The structure of the mining area is poorly known, due to infrequent outcrops. Strata in most mines are inclined at dips of 5° to 20°. There appear to be occasional high angle normal and reverse faults, but these are seldom of any significance in mining. The friable white silica zone mined in the Clear Creek chert is overlain by dark brownish-red chert and siliceous shale. Exploration drilling to date has not encountered the base of the microcrystalline tripolitic zone. Chemical content Leaching by ground water solution has removed virtually all the carbonate from the original siliceous limestone. Impurities encountered are red to yellow iron oxides, aluminous clays, and occasional black manganese oxides. The silica has no detectable formula water. A typical chemical analysis and physical properties are listed in Tables 1 and 2. Mineralogy and crystallinity Southern Illinois friable tripoli appears amorphous under the polarizing microscope. No crystal structure and only occasional, vague, patchy anisotropism is evident. However, the material gives an excellent quartz X-ray diffraction pattern. Thus, so far as X-ray mineralogy is concerned, the material is clearly

Jan 12, 1986

By J. B. Gwiazda

S.S. Peng The highlights of this paper are : • The shield supports with lemniscate links will exert a tangential force on the roof strata, causing unnecessary and harmful roof movements, thus damaging the roof structure. The principal source of the tangential force is the horizontal movement of the point O, the pivotal joint between the canopy and the caving shield. When the support height is changed during setting, the point O will move along a curve upward and rearward toward the gob. As a result, it produces a vertical movement, causing the roof to move up a distance of ?d, and a horizontal movement toward the gob, creating vertical fractures in the roof. • The major purpose for employing the lemniscate-link system is to protect the hydraulic legs against the action of the bending moments induced by the tangential force that causes deformation or destruction of the legs. • Based on these two points, the author suggested that, in order to eliminate the harmful effects of the lemniscate-link system and strengthen its advantageous features, a shield support with a new design of that mechanism has been developed. With the new design, the conventional rigid lemniscate links are replaced by hydraulic rams. When the pressure inside the rams reaches a certain level, the ram will yield.

Jan 12, 1986

By Lawrence Pelham

Introduction Fluorspar demand or consumption has been taken for granted in recent years, except for those that produce, market, or consume the commodity. Thinking used to be that as the steel and aluminum industry goes, so goes fluorspar. This is much less true today. The industry is becoming more dependent on chemical industries that feed primarily off hydrofluoric acid (HF) production. The US Bureau of Mines (USBM) published an assessment of world fluorspar reserves and resources. It created a foundation of valuable data that can be used to monitor fluorspar availability. Celebrations were held this year commemorating the 100th anniversary of the isolation of fluorine. The US and other countries continue to struggle with the controversy of ozone depletion and the effect of fluorocarbons on the atmosphere. Goals for the government's strategic stockpile are being reevaluated. The market for fluorine chemicals is changing and diversifying, making it a challenge to assess and forecast fluorspar/fluorine demand. US production and distribution Fluorspar Fluorspar shipments from US mines reached an all-time high near the end of World War II in 1944 at 375 kt (413,800 st). Eighty-two mines produced more than 453 t (500 st) each, and 10 mines produced more than 9 kt (10,000 st.) each. Illinois supplied 43% of total shipments. Consumption in 1944 was 372 kt (410,200 st), of which 56% was used by steel producers and 32% by HF producers. This was the last year that US production exceeded US consumption. The post-World War II record for fluorspar shipments occurred in 1951 at 315 kt (347,000 st) 58 mines produced more than 453 t (500 st) and eight mines produced more than 9 kt (10,000 st).

Jan 11, 1986

By S. G. Malghan

W.H. Moyer In reference to the section “Excess pyrite, ” the author discusses difficulties in separating chalcopyrite from pyrite. A technique does exist, and has been used commercially, for the flotation, separation, and recovery, if desired, of both chalcopyrite and pyrite. The method was described in Mining Engineering, March 1966, “An Improved Flotation Method for Cornwall Sulfides. ” While it was not used on a porphyry copper ore, it was used on a chalcopyrite - pyrite-containing ore and might well have application. Both sulfides, chalcopyrite and pyrite, were floated simultaneously using xanthate as a collector and MIBC frother to produce a cleaned sulfide product. This concentrate was then thickened and conditioned at high solids with aeration and lime to a pH above 12.0 for a period of 1.5 to 2 hr. A supplemental suppressant, such as cyanide, could be used if necessary to sharpen the subsequent separation; however, it was not required at Cornwall. The conditioned sulfides were floated with two stages of cleaning to produce a highgrade chalcopyrite froth product and a pyrite sink product barren in copper. The separation at Cornwall gave a 28.6q. Cu grade chalcopyrite product and a pyrite analyzing 0.0670 Cu. A desliming step on the raw ore initially used at Cornwall for a period of time was later discontinued. Advantages to be found in using a strong sulfide collector, such as a xanthate, is the rapid flotation it induces and, generally, an ability to float at a coarser grind. A following regrind circuit could be used to give maximum liberation and produce a clean sulfide concentration. The “differential” separation is concerned primarily with grade and with assuring a barren pyrite for maximum recovery.

Jan 11, 1986

By J. B. Hiskey

Introduction The domestic copper industry has suffered for some time from declining ore grades; rising costs associated with conventional mining, milling, smelting and refining; tightening environmental restrictions; and shrinking markets due to materials substitution. Ore grade depletion is a major problem. It is illustrated by the decline in average copper yield shown in Fig. 1. From 1979, copper grade has dropped 60%, lowering yields from 1.2% to 0.49% (Sousa, 1981). The US copper industry is taking drastic measures to enhance its competitive posture and to guarantee its future worldwide position. In a sense, this involves a general restructuring of the processing philosophy of US copper producers. Judging from recent trends, this restructuring relates to the application of solution mining techniques to the treatment of copper ores and low grade waste material. Recovering copper by leaching techniques has been practiced for centuries. As early as the mid-16th century, some Hungarian copper mines were recycling copper bearing leach solutions through waste heaps (Nash, 1912). In the US, recovering copper from dilute mine waters has been carried out for more than 100 years. There are, however, certain technical innovations that have elevated the importance of solution mining as a process for copper recovery. These include: • Advances in solvent extract-electrowinning (SX-EW); • Developments in acid and acid ferric cure processes for oxide and mixed oxide sulfide ores; • Improvements in heap and dump construction; and • Advances in in situ leaching technology. Generally, solution mining enjoys certain intrinsic advantages over conventional mining and milling. Compared to conventional milling, the combined capital and operating costs of leaching facilities are normally lower, start-up times are faster, and the leaching operation usually has less impact on the environment. Furthermore, solution mining represents an expedient way of extracting metals from small shallow deposits, and is particularly suited to the treatment of low grade sources. And, in the case of waste dump leaching, there is a tremendous resource in-place that has the mining cost already off the books. This paper updates current US copper solution mining trends. Solution mining systems Schlitt (1980) summarized the characteristics of the principle leach systems in the US. He pointed out that the wide variations in copper deposits make solution mining a more site-specific activity than conventional mining and milling. Heap leaching is usually selected for moderate to high grade ores having a copper mineralization that is amenable to acid leaching (oxides, silicates and certain sulfides). Higher grade ores mined specifically for heap leaching are usually treated to optimize copper extraction, therefore, crushing and some pretreatment (acid curing) may be justified. On the other hand, dump leaching practice usually involves leaching low grade waste rock placed on large dumps as dictated by topography and haulage costs. Mine waste is therefore treated in an uncrushed run-of-mine condition. Finally, in situ leaching is appropriate for either very deep ore bodies or for recovering copper from low grade rock left behind in abandoned pit walls, stopes, and subsidence zones from earlier mining activities. True in-situ solution mining should be defined as the leaching of ore in its original geologic setting. Heap leaching Heap leaching has a long history as a hydrometallurgical process for copper recovery. The methodology of heap leaching established at Rio Tinto in Spain more than 300 years ago is basically the same as that used today. Processing concepts, such as solution management (sprinklers and leach/rest cycles) and copper

Jan 11, 1986

By Syd S. Peng

Introduction The most commonly used ground control design in coal mining is the determination of pillar size by various pillar strength formulae for mine layout (Bieniawski, 1983; Holland, 1973; Hustrulid, 1976; Obert and Duvall 1967; Wilson and Ashwin, 1972). However, there have been few documented case histories concerning the validity of those strength formulae. Recently there are indications that those strength formulae tend to overdesign considerably. From a safety point of view, there is nothing wrong with overdesign. But an overdesign with large chain pillars reduces the flexibility of machine movement and decreases production. It also reduces coal recovery and, thus, shortens the reserve's life. This is especially true for small operators who own a limited reserve. Those small mines, through past practices or out of necessity, experiment with smaller pillars than those predicted by the strength formulae. Conclusive ev¬idence has been collected that points to the need of a new concept of coal pillar design. Two case studies will be cited to illustrate this point. 3-D US pillar strength formulae Figure 1 shows a sectional layout of a one-unit mine. The seam was about 1.3 m (50 in.) thick with a uniform cover of 64 m (210 ft). A house was located near the center of the section. The homeowner demanded that no surface subsidence and domestic well dewatering should ever be allowed as a result of underground room-and-pillar mining. Physical properties - strength and Young's modulus - of the rock strata including coal were obtained by testing small samples in the laboratory. The uniaxial compressive strength of the coal was 3660 kPa (823 psi). Using various pillar strength formulae, the homeowner recommended that square chain pillars be at least 11 m (35 ft) wide. The coal company, however, desired to pursue optimum coal recovery in the safest and most economical manner. They engaged in three-dimensional finite element analysis of the section taking into account the complete stratigraphic sequence of the overburden. Results indicated that the section can be developed by 5.5 m (18 ft) wide entries at 21 m (70 ft) center-to-center. In retreat, the pillars can be splitted along both centerlines leaving a stump pillar of 5.5 m (18 ft) square at each of the four corners. The 5.5 m (18 ft) square stump pillars still maintained a safety factor of 1.5, although the safety factors calculated by various pillar strength

Jan 11, 1986

By A. G. Sciulli, G. P. Ballock, K. K. Wu

Waste disposal underground is the most effective method when viewed from an environmental aspect. As costs of other methods of disposal continue to escalate, the cost effectiveness of underground disposal during the mining cycle becomes increasingly attractive. The use of concrete for sealing abandoned mines has proved to be suspect when associated with a relatively incompetent adjacent strata due to its shrinkage characteristics. Complete failure of the plug has been known to occur in some instances. A more effective, cost competitive seal can be obtained using pneumatically stowed material and incorporating a polyurethane grout, which has greater adhesion characteristics than concrete. Injection of the surrounding strata and of the stowed material helps to ensure an impervious barrier and prevent the plug from moving when subject to high pressure heads. When properly placed, impervious seals cause the old workings to flood, raise the mine pool level, and thus eventually eliminate the potential for oxidation, which is the primary cause of acidity in the water.

Jan 11, 1986

By W. M. Luff

Introduction The Brunswick No. 12 mine is located about 27 km (17 miles) southwest of Bathurst, New Brunswick. The deposit is a stratiform massive sulfide deposit occurring in a series of sedimentary and volcanic rocks of the Tetagouche Group. These rocks are generally considered to be Ordovician in age. Ore reserves at Brunswick No. 12 are 104 Mt (115 million st) grading 3.72% lead (Pb), 9.01% zinc (Zn), 0.32% copper (Cu), and 97.9 g/t (2.8 oz per st) silver (Ag). Production in 1984 was 3.5 Mt (4 million st) at 3.57% Pb, 8.91% Zn, 0.32% Cu, and 100 g/t (2.9 oz per st) Ag. Total payable silver production that year was 211 t (6.8 million oz). This made Brunswick one of the largest silver producers in Canada. Bathurst-Newcastle area and Brunswick Mining Interest in the Bathurst area was aroused in 1897 when W. Hussey discovered an iron formation at Austin Brook, just south of Brunswick No. 6 deposit. This area, and the current No. 6 area, was explored from 1897 to 1907. Iron ore was mined at Austin Brook from 1907 to 1913 and from 1942 to 1943. Any occurrence of sulfides was overlooked until 1952. In 1952, M. J. Boylen Enterprises investigated the small sulfide body associated with the Austin Brook iron formation. Other electromagnetic anomalies associated with the iron formation were investigated. The No. 6 deposit was found late in 1952, when hole B12 intersected 97.5 m (320 ft) of massive sulfides. The No. 12 deposit, 10 km (6 miles) to the north, was found in 1953. More than 30 massive sulfide deposits have since been identified in the Bathrust area. In the 1950s, St. Joe Lead Co. did extensive test work on the No. 6 and No. 12 deposits, including shafts at No. 12, lateral development at No. 12, and testing at both deposits.

Jan 10, 1986

By Robert A. Jankowski, J. Harrison Daniel, Fred N. Kissell

Introduction Longwall coal mining in the US has steadily increased since its introduction in 1960. There are now 108 systems operating. (Sprouls, 1986). Advantages of longwall mining include higher productivity, improved safety, and greater coal recovery. Respirable dust, however, is one of the more serious obstacles to realizing the full potential of longwall mining. Early in 1980, the US Bureau of Mines (USBM) initiated extensive in-house and contract research to identify and evaluate effective methods of controlling respirable dust exposure of longwall face personnel (Niewiadomski, Jankowski, and Kissell, 1981). It was found that several control procedures must be combined to assure compliance, using control techniques appropriate to the dust source. These techniques can typically be grouped into the following categories: reduction of dust generation; dilution of dust; capture or knockdown of dust; and dust avoidance. Control techniques can be divided into basic and effective dust control techniques; additional necessary controls; and optional supplementary controls. This information should allow longwall mine operators to select control methods applicable to their particular operations and to develop a dust control plan that meets Federal dust standards. Basic and effective dust control techniques There are three basic dust control techniques (Taylor and Jankowski, 1982) that should be implemented on all longwall mining operations. • A passive barrier or external water spray system design that confines shearer generated dust near the face and away from the operators. • Large quantities of water through the sprays, particularly the drum sprays, to aid in dust knockdown and prevent dust becoming airborne. • A cutting sequence that allows the shearer operators to work primarily on the intake-air side of the lead cutting drum. Shearer clearer There are some poorly designed shearer-mounted spray systems with nozzles directed upwind at the cutting and loading zone of the intake-side drum. These actually carry dust away from the face and upstream of the drum. Where, it mixes with the clean intake air (Fig. la,) and is carried out into the walkway over the shearer operators. The Bureau developed a novel shearer spray system, called the shearer clearer. It takes advantage of the air-moving capabili-

Jan 10, 1986

By V. Rajaram

R.K. Leininger, N.R. Hasenmueller, and D.D. Carr The content of the paper by V. Rajaram does not fulfill the expectations of the title. But Rajaram submitted the article in October 1983, and he could not have foreseen the numerous developments that would occur before his paper was published more than two years later. Nevertheless, Rajaram failed to mention the interest in oil shale development in southern Indiana beginning in late 1979 and continuing through the present as a result of special financial encouragement of three oil shale projects in May and August 1984 by the Indiana Energy Development Board and the Indiana Corp. for Science and Technology. A session at the 1985 Eastern Oil Shale Symposium in Lexington, KY, Nov. 18-20, 1985, gave the current status of oil shale developments in the eastern United States. Speakers at this session reported on the three Indiana projects. First, Gary D. Aho, Cliffs Engineering Inc., spoke on a feasibility study by Cliffs Engineering and Allis Chalmers for a site-specific pilot plant using the Allis Chalmers process. The plant site is in Clark County, IN, on property of the Midwest Energy Resources Co. The project is funded by $240,908 each from the Indiana Energy Development Board/Corp. for Science and Technology (EDB/CST) and the US Department of Energy (DOE) and $120,454 from the corporate sponsors (total $602,270). Completion is scheduled for mid-1986. Next, Edwin M. Piper, Stone and Webster Engineering Corp., discussed the American Syn-Crude/Indiana Oil Shale Project. This effort followed the completion of two smaller projects funded by the Indiana EDB/ CST: "Testing of Indiana Oil Shale in a Petrosix Pilot Plant" (total funding by EDB/CST at $50,000) and "Assessment of the Petrosix Process for an Indiana Shale Oil Plant" (total funding by EDB/CST at $100,000). The status at the time of the report at the Oil Shale Symposium was that a request for an extension of time for securing nonfederal support had been submitted to the US Synfuels Corp. The proposal included building an 11-m (36-ft) diam retort in southeastern Indiana to process the New Albany Shale and to produce 366 m3 (2300 bbl) of shale oil per day by the PETROSIX process. The Indiana EDB/CST had contracted with Stone and Webster at $401,100 from EDB/CST and $245,835 from Stone and Webster for activities to advance the project in the negotiations with Synfuels Corp. At the time of the report, this project was the only eastern oil shale proposal that was still on the agenda of the Synfuels Corp. As a result of Congressional action in late December 1985, federal support from the Synfuels Corp. is no longer possible. Finally, Victor H. Carr, Eastern Shale Research Corp., described his firm's project, which is jointly supported by DOE ($227,749) and the Indiana EDB/ CST ($73,850) and is entitled "Feasibility Study to Determine Suitability of an In-Situ Process to Recover Hydrocarbons from Eastern Shale." An area in Scott County, IN, had been chosen, but not a specific 9 x 15-m (30 x 50-ft) site. One burn of a small in situ retort is contemplated as part of the project. Besides these three projects, several reports on shale research were presented. Joseph Damukaitis, American Syn-Crude Corp., reported that a pilot plant using the hydrogenation-extraction (H-E) process (described at the 1984 Eastern Oil Shale Symposium) was 94% built and would go through shakedown with oil shale but would then shift mechanical devices to process coal mine waste. Current status of research on the HYTORT process was then presented by Raymond C. Rex, Jr. Oil shale beneficiation research at the University of Alabama/Minerals Research Institute was reported by R. Bruce Tippin. Scott D. Carter discussed continued research on fluidized-bed retort¬ing of shale at the Kentucky Center for Energy Research Laboratory. Carl E. Roosmagi of DOE, Morgantown, reported on the oil shale research at the Morgantown Energy Technology Center (METC) laboratory. Henry J. Gomberg, Ann Arbor Nuclear Inc., discussed "Radiation Combined with Donor Solvents for Extraction and Up-Grading of Kerogen." Aurora M. Rubel and coauthor Eileen Davis presented results of research at the Kentucky Center for Energy Research Laboratory under the title "The Effect of Shale Particle Size on the Products from the Bench Scale Fixed Bed Steam Pyrolysis of Kentucky Sunbury Shale." Lastly, Maria Rockwell, Technical University of Nova Scotia, presented "Processing and Up-Grading of Low Grade Nova Scotia Oil Shale for Potential Use." A review of shale oil prospects by Gerald Parkinson in Chemical Engineering for Feb. 3, 1986, covers both western and eastern projects and includes a report that the US DOE is funding a few relatively small projects; most of the fiscal 1986 budget of $12 million for shale oil is for fundamental research. Projects include a $3.2 million three-year contract with Hycrude Corp. (Chicago) for development of the HYTORT process and a $1.2 million three-year contract with the University of Alabama's Mineral Resources Institute on beneficiation of eastern oil shale by froth flotation. The report incorrectly states that current projects include a total investment of $466 million in a pair of 18-month technical and economic feasibility studies for proposed projects in southern Indiana. The correct figure is $468,657 [$240,908 for the Cliffs Engineering/

Jan 10, 1986

By R. K. Singhal, H. Sahay

Introduction In terms of the world's recoverable coal reserves, Canada ranks 10th, providing 0.6% of the world's total. The US is first with 27.8% and the Soviet Union follows with 24.1%. Figure 1 shows the principal coal mines in Canada. Several provinces contain substantial coal deposits, but the largest deposits are found in the western provinces of Alberta, British Columbia, and Saskatchewan. Conservatively, 90% of the lignitic, 80% of high volatile bituminous, and nearly 100% of subbituminous, and low and medium volatile bituminous coals are located in these three provinces. During the past decade, Canadian coal production of all types increased from 18.8 Mt (2 million st) in 1972, to 25 Mt (28 million st) in 1975, to 61 Mt (66 million st) in 1984. About 38 Mt (43 million st) of this output was classified at thermal and 23 Mt (25 million st) as coking and metallurgical coal (Table 1). The three western provinces currently produce 57 Mt (62 million st) of coal - 95% of Canada's total. The remaining 3.6 Mt (4 million st) (5%) is produced in Nova Scotia and New Brunswick on the east coast. The majority of western coal (98%) is produced by surface mining methods. On the east coast, though, mining is predominantly underground and only 15% is produced on the surface. In 1985, Canada exported close to 28 Mt (30 million st) of metallurgical coal and thermal coal. During the same period, 15 Mt (16 million st) of coal were imported, mostly from western US for use in Ontario and Quebec. Almost 60% of the imported coal was used for thermal applications. The remaining 40% was metallurgical coal for the steel industry. It is generally cheaper to import coal into central and eastern Canada from the eastern US to augment local supplies rather than from western Canada. This is mainly because of high transportation costs. Mining methods The suitability of mining methods in a particular area depends largely on the geological setting. Geology, therefore, simplifies or adds complexities to mining operations, directly affecting the economics. It also influences the flexibility required from the min-

Jan 9, 1986

By K. Hanumantha Rao, A. Nayak, K. S. Narasimhan, S. N. Mahapatra

B.A. Hancock It is not at all surprising that causticized potato and potato derived amylopectin starch solutions performed much better than their parent starches. Some preparation is required to rupture the starch granules to effect the polymeric adsorption and interparticle bridging necessary for selective flocculation. In laboratory work comparing the deslime performance of causticized and autoclave cooked laboratory corn starch solution preparations, it was found that higher deslime weight rejections, with attendant proportionally greater iron unit losses, occurred with the causticized starch. These results may be specific to the ore involved but they do suggest that cooking and causticizing cause different starch granule rupture and/or starch breakdown, which have an effect on desliming response. I calculated from the data in the article that the slimes product grades were high - 24.3% and 20.3% Fe when 53.7% and 54.5% Fe concentrate products were obtained, respectively, in Table 4, and 21% Fe with a 62.6% Fe concentrate in Table 5 - using the natural tailings sample, which had a head of 34.3% Fe. It may be advisable for the authors to consider different starch preparations in future investigations. The combination of upgrading and selectivity results presented in Table 4 are not as good as the authors suggest. The authors' claim that a system has been developed to produce saleable concentrates from the Kudremukh tailings is quite disconcerting. There are many hurdles yet to be crossed before commercial application of selective flocculation becomes possible because differences between the very small-scale laboratory tests conducted and commercial application are rather large. Among the many differences are varying circuit feed grades that will occur from use of tailings, the apparent face that much lower tailings grades will be encountered in practice (it is much easier to achieve a high concentrate grade with reasonable recoveries using 34.3% Fe tailings as in the study rather than 25.3% Fe tailings grades that the plant apparently averages), the hydraulic nature of the thickeners used in operations compared to the static system used in laboratory tests, the different size distributions that will be obtained from a plant closed grinding-classification circuit, and differences in water used in a plant operation and the laboratory. The authors wrote that it was necessary to overgrind to be sure that the coarse gangue would not settle with the iron oxide floccules. This situation is likely to be exaggerated in commercial operations where it is assumed cyclones would be used for classification. Because cyclone classification is greatly influenced by particle densities, there will probably be an even greater difference in size between the iron and gangue particles in the plant, which would make the gangue slightly coarser still in relation to the iron. This would make the selective flocculation-desliming separations using the procedure employed by the authors even more difficult and, using the dispersant system the authors employed, greater overgrinding would be required. To grind finer to minimize the coarse gangue in the flocculated iron oxides is quite inefficient and appears not to broach the problem. The actual problem appears to be insufficient dispersion of the ground pulp. In this situation, addition of a dispersant would likely be required to attain a sufficiently high pulp dispersion level to efficiently effect a selective flocculation-desliming separation. Although the very coarse particles would still have a tendency to settle with the floccules, it probably would be found unnecessary to overgrind as much as indicated. Use of an optimum combination of dispersant and pH modifying reagents may also significantly improve the selectivity of desliming. Additionally, although it is possible that sufficient dispersion may be obtained by pH control alone in some situations, it is quite probable that added dispersity was obtained in the reported work from using distilled water. It is research experience that distilled water enhances dispersion. In commercial operations it may not be expected that sufficient dispersion will be obtained by pH control alone, unless the water used in the process is by nature quite dispersive. Overall, a change in the Kudremukh tailings dispersant scheme appears necessary where a dispersant is used in conjunction with a pH modifying reagent. With this change, different dispersion-flocculation responses will result that would have to be further evaluated. Therefore, it is still an open question whether an efficient and effective selective flocculation separation using Kudremukh tailings may be obtained that will produce saleable concentrates.

Jan 9, 1986

By A. L. Crandell

History The Cortez Mining District has a long history of mining activity dating back to 1862. More recently, Cortez Gold Mines had operated a cyanide leach, counter current decantation (CCD) recovery gold mill from January 1969 to February 1976. This type of plant has been standard in the US for treating higher grade, finely disseminated gold in Carlin type ores. The mill was initially operated from 1969 to 1973 at 1.5 kt/d (1650 stpd) on oxidized ore from the Cortez ore body. In 1973, the mill was operated at 2 kt/d (2200 stpd) on softer, less oxidized, high clay ore from Gold Acres, located 13 km (8 miles) west of Cortez, NV. In February 1976, the oxidized portion of the ore body at Gold Acres was depleted and the mill ceased operations. Successful heap leaching of the lower grade portion of the ores was practiced at both Cortez and Gold Acres to supplement the milling operations. Valuable experience was gained in the operation of these two heap leaching programs (Duncan and Smolik, 1977). The ore heaps at Gold Acres and Cortez were releached with a limited crew from 1976 to 1979. The surge in the price of gold in late 1979 sparked an interest to incorporate new techniques into treating lower grade and less oxidized carbonaceous ores at Cortez. In early 1980, heap leaching was started on low grade oxidized dump material left from the earlier Cortez mining operation. By June of 1980, testwork was completed on the London Extension Dump, a low grade, mildly carbonaceous dump at Gold Acres. It was decided to convert the mill at Cortez from the conventional cyanide leach CCD recovery plant to a carbon-in-leach (CIL) recovery system. In May 1983, ore processing from the Horse Canyon deposit began. The Horse Canyon ore body contains both oxide and carbonaceous ores and is located 22.5 km (14 miles) from the Cortez mill. The Cortez CIL circuit was expanded to handle a blend of oxide and carbonaceous ores from Horse Canyon. Ore selection In the last quarter 1979, gold prices were increasing to a point where restarting the Cortez mill, idled since 1976, appeared imminent. Several previously explored geologic reserves at various stages of development were available as possible sources of feed for the mill. Initial metallurgical testing of dumps and tailings from earlier operations was completed, and previously uneconomic deposits were explored. It was decided to pursue sources of ore with the least lead time to production and those with the least capital cost requirements. A master plan was developed to bring the four profitable ore resources into production. The plan called for: • Heap leaching of waste dumps from the Cortez mining operation; • Restarting the mill to handle the waste dumps from the London

Jan 9, 1986

By D. R. Nagaraj

Editor&apos;s note: Minerals concentration is a disparate area, one involving many disciplines. Flotation, flocculation, reagents, microbial processes, gravity and magnetic separation, and the areas of coal and oil shale - concentration is relevant to all of these. In his article, Nagaraj reviews the latest developments and findings in concentration that have ocurred during the past year. Flotation The Monograph Series "Development in Mineral Processing," (Elsevier, 1985) was devoted to "Flotation of Sulfide Minerals." This volume has 26 articles covering the aspects of flotation chemistry, process development, plant design and equipment, and modeling and process control. The numerous articles on sulfide flotation had one recurring theme - the role of electrochemical processes and redox potentials in flotation and depression. Significant advances have been made in this area. A differential desorption-reflotation process was developed for the beneficiation of calcitic phosphate pebble (Hsieh and Lehr, Industrial Engineering and Chemistry, 1985). The carbonate and phosphate minerals were first separated from silica using fatty acid and the modifiers diphosphonic acid and HF. The collector on phosphate was then desorbed using sulfuric acid, and calcite was selectively floated without additional collector. A two-stage conditioning - initially at pH 10, followed by that at pH < 4.5 - with sodium oleate was developed to improve the separation between apatite and dolomite (Moudgil and Chanchani, Minerals and Metallurgical Processing, MMP, 1985). Separation of Mussorie rock phosphate in India from gangue calcite, dolomite, silicates, pyrite, and carbonaceous impurities was achieved with a three-stage flotation only on -75 µm (-200 mesh) fraction (Vaman Rao et al., International Journal of Mineral Processing, IJMP, 1985). The separation of fine SiC whiskers from other products of coked rice hulls by immiscible liquid and froth flotation was achieved using M1BC, pine oil, and mineral spirits (Parekh and Goldberger, MMP, 1985). Flotation was more economical. The carrier flotation separation of hematite fines from quartz fines was effected using coarse hematite as a carrier and oleic acid and ferric chloride (Cristoveanu and Meech, CIM Bulletin, 1985). Potash recovery from brines by laboratory-pan solar evaporation and amine flotation was discussed by Foot and Juiatt (Salts and Brines &apos;85, Proceedings, 1985). Flotation in a 45-kg/h (99-lb per hour) unit recovered more than 95% of the potash in schoenite and sylvite concentrates containing 28% and 62% K2O, respectively. The effects of fresh concrete into cement and aggregate by a single-step flotation process in aqueous suspension was described (Naegle, Chemical Engineering Science, 1985). The physico-chemical principles for the selective separation of proteins from mixtures using flotation were discussed by Ostermaier and Dobias (Colloids & Surfactants, 1985). Selective flocculation of hematite from silica was studied by Bagster and Mcllvenny (IJMP, 1985). Selective flocculation was more favorable with more highly charged anionic polyacrylamides and with high molecular weight. The cationic and non-ionic polymers flocculated both minerals. In minerals mixtures, there appeared to be very little selectivity unless a combination of dispersant and electrolyte was used. The effect of dissolved minerals species, such as Cu, Fe, and Ni on flocculation of chalcopyrite, pentlandite, millerite, pyrrhotite, and covellite was studied by Acar and Somasundaran (SME-AIME, 1985). Reagents There were some major advances made in reagent development for the minerals processing industry. Cyanamid introduced several classes of collectors, one of which - the diaryl monothiophosphates as acid circuit col¬lectors of sulfides and oxidesulfide ores - was commercialized in 1985. The monothiophosphates are more stable and powerful collectors than the xanthates and dithiophosphates. Hydrolyzed dithiophosphate containing 80% monothio and 20% dithiophosphate was found to give optimum recovery of copper from crude cementation precipitate and separation of Cu and Ni sulfides from matte (Akimova et al., Obogashch, Tonkov, Rud, OTR, 1985). A new class of O-alkyl-Nallyl

Jan 9, 1986

By V. Rajaram

R.C. Rex, Jr. V. Rajaram&apos;s paper does not accurately reflect the status of Eastern oil shale development as it stands today. Today, there are numerous firms and research organizations pursuing development of the Eastern oil shale resource. Details of the current projects can be found in the symposium proceedings for the 1985 Eastern Oil Shale Symposium held in Lexington, KY, in November 1985. Additionally, the comparison of the Paraho and HYTORT processes is invalid because the HYTORT process design used it not an accurate representation of the HYTORT process. I will briefly address this point. It is disconcerting to see the author reference hydroretorting work conducted at the Institute of Gas Technology (IGT) (circa 1978) in the introduction, which indicates hydroretorting (the HYTORT process concept) yields of up to 250% of the Fischer Assay oil yield, and then turn around and base the remainder of the paper on the 1981 "Synthetic Fuels From Eastern Oil Shale Report," which depicts the HYTORT process as yielding only 115% of Fischer Assay. Indeed, work conducted at IGT to 1981 indicates commercial oil yields for the HYTORT process would be in the range of 200% to 210% of Fischer Assay. A more thorough literature search by the author would have yielded a 1981 publication by Schora et al., entitled "Update on the HYTORT Process" presented at the Symposium on Synthetic Fuels From Oil Shale in Nashville, TN, October 1981. This paper thoroughly discusses the shortcomings of the referenced study with respect to the HYTORT process. For the interested reader, I can also recommend a June 1984 report entitled "An Economic Comparison of Five Process Concepts for Using Eastern Oil Shale," published by Los Alomos National Laboratory (LA-10106-MS, Parkinson et al., NTIS No. DE84-015981). This study applies a realistic HYTORT design (circa 1982) and indicates that HYTORT is the most econom¬ical method of processing Eastern oil shales (about $20/bbl less than Paraho). Additionally, completion of the HYTORT feasibility study in June 1983 by Phillips Petroleum Co., Bechtel Group Inc., HYCRUDE Corp., and IGT has resulted in development of designs that will make the use of HYTORT on Eastern oil shales even more economic. HYCRUDE Corp., a second-tier subsidiary of IGT, is involved in commercialization and licensing of the HYTORT process and is active in efforts to develop the Eastern oil shale resource. As such, I am disappointed to see outdated material that has the potential for adversely affecting development of Eastern oil shale resources published in MINING ENGINEERING. I can only hope that this response will serve to mitigate any negative opinions that may have been formed. Today, activity in commercializing Eastern oil shales dominates the oil shale scene in the US and indicates a realization of the potential for these shales. Reply by V. Rajaram I appreciate the updates on the HYTORT process provided by R.C. Rex, Jr. I hope that Hycrude&apos;s work will lead to the funding of a pilot plant to demonstrate that yields of greater than 200% of Fischer Assay can be achieved in a commercial plant using eastern US oil shales.

Jan 9, 1986

By C. D. Elifrits, N. B. Aughenbaugh

Introduction Federal and state laws enacted to regulate coal mining and the accompanying public concern about the adverse effects that mining might have on land use have focused much attention on the subsidence of the land surfaces over underground mines. Some questions asked are: • Will abandoned mines subside in the future even though they have exhibited no surface adjustment to date? • If so, when will subsidence take place? • Can a specific surface area be affected by subsidence more than once? • What surface form will the subsidence feature take? • Is there any way to predict when subsidence will occur and end? An attempt is made here to address these questions insofar as it is possible with available data and experience of evaluation of subcritical extraction conditions. Mining that results in subsidence features can be classified as supercritical, subcritical, and critical with respect to the minimum span lengths of the worked-out rooms (Fig. 1). A complete basin will not develop fully immediately over subcritical conditions because the span lengths of the entries are less than the critical dimension (Wardell, 1983). Passageways, entries, and room-and-pillar panels are all included in the subcritical category. Because of the short span lengths, subcritical mining conditions tend to be stable over the life of the mine unless adverse geologic conditions exist (Gray et al., 1974; Dunrud, 1976). Various roof control and support procedures are employed to maintain entry stability during mining. Little information exists concerning the time factor of subsidence over areas of subcritical extraction. Data that have been collected relate to longwall mines, whereby the cavity sizes are designed large enough to induce caving behind the working faces so that complete subsidence at the surface occurs within a short period of time. The UK&apos;s National Coal Board (1975) has reduced the enormous amount of data it has collected to empirical correlations and has published these correlations in its Subsidence Engineers Handbook. The handbook states that "None of the foregoing remarks applies to pillar and stall working," with respect to the section discussing the time factor of subsidence caused by longwall mining (National Coal Board, 1975, p. 40). To understand and evaluate how time relates to subsidence above room-and-pillar mines, one must first recognize how mine cavities deteriorate and fail. The propagation of failure to the surface must then be studied, and, finally, the failures that are manifest at the surface in the form of depressions accompanied by both compressional and tensional movements must be investigated. Failure of mine cavities Any cavity beneath the earth&apos;s surface represents a condition of nonequilibrium and tends to close with time. Failure of cavities in a subcritical mine can be classified as failure of the roof, failure of the floor, failure of the supports (pillars, cribs), and elastic rebound of the rock. One mode of failure may dominate, but generally combinations of the four will occur. Also, initial instability of geologic materials and structural features and in situ stress may be unknown factors that can contribute to cavity failure. Roof failure is the most obvious opening stability problem in most operating mines. The bridging capabilities of the roof rock are dependent on the rock&apos;s material strength, geologic discontinuities, and reaction to humidity changes in the air of the mine. This type of failure may vary in magnitude from gradual flaking to massive roof falls. Mine floor instability can be manifest in the form of a general heave over a broad area or of individual pillar punching. The weaker and thicker the immediate floor material, the more prone a mine is to experience floor problems. Water seeping into an area can soften the underclay, reduce its bearing capacity, and precipitate failure of support for the overburden.

Jan 8, 1986

By Gerald M. Ellis

ntroduction As the gold mining industry continues to develop in the Dominican Republic and Haiti, more deposits will undoubtedly be discovered that will be amenable to development. The Pueblo Viejo gold mine is currently in operation in the Dominican Republic. It has become an important component of that country&apos;s economy. Pueblo Viejo output is important to the government as a foreign currency earner. It is also said to be among the world&apos;s more successful gold mines. But, while Pueblo Viejo is successful, no other major gold discoveries have been made in recent years. Therefore, alluvial deposits of gold have caught the attention of geologists on the Island of Hispaniola. In the past four years, exploration and investigation of alluvial deposits have been initiated in both the Dominican Republic and Haiti. This type of activity is expected to continue over the next few years. Alluvial gold mining is rudimentary gold washing, done mostly by farmers. It is an old tradition, dating back at least 100 years, and is done by small, independent, or community farming villages in the mountains and plains of the Dominican Republic. Investigations in the DR Recently, alluvial gold investigations have been undertaken in well known locations of the Dominican Republic. One is in the Villa Altagracia area. Here, the terraces and gravels of the Haina River basin were examined, test-pitted, and assayed by Minera Altagracia, a local Dominican mining group. Results led to a new investigation by the state-owned Rosario Dominicana (RD), the company that mines the gold at Pueblo Viejo. Rosario tested areas previously investigated by Minera Altagracia. RD could not define the type of concentration that could justify an industrial-scale type of exploitation. The Haina River placers are located in the flats and valleys of Central Dominican Republic. They are classified as small-scale type of operations where local cooperatives can do the traditional gold washing. It is not now considered a dredgable project for a mining company. The next alluvial gold investigation was undertaken by the Dominican Department of Mines and others in an area of some 700 km2 (270 sq miles) in the northeast near Miches. Farmers have been washing gold in streams and excavating pits in various areas, particularly La Mina and have made a living. Government geologists panned several hundred streams and river tributaries. They also washed material from deep excavations. The results, however, were negative insofar as locating alluvial gold concentrations suitable for large-scale or industrial-scale exploitation. The most favorable locations were the terraces along the Jayán and Cuarón Rivers in the area of Miches. But the grade od concentration is poor. Actually, only traces of gold were found in these

Jan 8, 1986