Search Documents

By F. E. Pawlek

Copper concentrates can be comminuted in an attritor within a very short time to a particle size of 1 - 0.1 um. Addition of sodium hydroxide enhances commninution. The concentrates, suspended only in water, can be completely dissolved under oxygen over pressure and temperatures ranging from 120 - 180°C (248 - 356°F), because the surface area of the concentrate has been significantly increased by the ultra fine grinding. Addi¬tion of a small amount of silver ions to leach liquor containing sulfuric acid acted as a catalyst for the dissolution of chalcopyrite, because of the formation of local elements on the surface. The addition of C1- ions was also catalytically very effective. With this method it is possible to completely dissolve the copper contained in chalcopyrite concentrates after ultra fine grinding within 30 min at 110°C (230°F) under an oxygen partial pressure of 1 atm (14 psig). Several flowsheets demonstrate the industrial feasibility.

Jan 1, 1976

By Kailash Narayan

Atlantic Richfield Company began operating a commercial uranium solution mining project in April, 1975 in Live Oak County, 10 miles southwest of the town of George West (Figure 1). The operation is designed to produce 250,000 pounds of U303 per year and is jointly owned by Atlantic Richfield Company, Dalco and U. S. Steel. The proprietary process employed in this operation consists of injecting an alkaline leaching solution into the ore-bearing formation through a pattern of injection wells. The uranium-bearing solution is then removed through producing wells using submersible pumps in central production wells. The produced solution is pumped through a pipeline to the plant area where the uranium is extracted. The barren solution is refurbished and returned to the pattern area for reinjection. The injection field and plant site are located in the drainage of Spring Creek, a small intermittent stream tributary to the Nueces River. Land use in the area is primarily agricultural -- cattle ranching and farming -- but several producing gas fields are present in the areas as well.

Jan 1, 1976

By R. A. Guidotti

The Federal Bureau of Mines studied the thermal decomposition of the nitrides of the Group V A metals (CbN, Ta3N5, and VN) with the goal of preparing high-purity metals. Solid-state purification and arc and electron-beam melting were examined. Electron-beam melting proved to be the most efficient means of nitrogen and oxygen removal. Columbium and tantalum of 99.9-pct purity were prepared by this technique; in the case of vanadium, however, nitrogen could not be reduced below the solid-solution level.

Jan 1, 1976

By John C. Franklin

A radon survey was made at five different locations in the Twilight mine near Uravan, Colo., to determine chances in radon concentrations in an inactive mine. Data were taken with a continuous monitor set for 1-minute count periods and readings were taken 24 hours a day for 9 days. The 12 readings per station each hour accurately showed how t w radon concentration varied during the day. From the data collected from the intake portal to station 4, the average radon emanation rate was estimated at 18 picocuries per second per square foot (pCi/sec/ft3). A daily cycle in radon concentration was detected that apparently was related to barometric pressure change.

Jan 1, 1976

To provide a faster means to open and close permissible electrical enclosures. Approach: An inspection and repair port with a multiple start thread which provides a permissible joint has been developed and tested. How It Works: The inspection port screw thread has a large pitch, which allows the port cover to be rotated from fully closed 10 fully open, in less than one revolution. See Figure 1. When the cover is screwed back in. it be-gins engagement within one eighth turn because the screw design has eight separate threads. To prevent Its unauthorized opening, the port cover has a bolt guarded by ears which accept a padlock. The threads are designed to tolerate a reasonable amount of grit without binding, although wipe-off cleanness is necessary during reassembly. Enclosures of two different sizes each fitted with the quick-access port, were

Jan 1, 1976

By James H. Allen

The air lift method of pumping water is discussed. A review of several large diameter hole drilling projects that used reverse circulation air lift techniques is presented. Proposals are made for reverse circulation flow rates for slow rates of bit penetration (5 fph) as well as reverse circulation flow rates for faster rates of bit penetration (10 fph). The Ingersoll- Rand air lift equations are modified for rotary drilling operations. By establishing minimum reverse circulation flow rates and utilizing the modified air lift equations, the establishment of minimum circulation equipment sizes and a more definitive requirement for improved bit body design should result.

Jan 1, 1976

By James A. Powell

The theory of electromagnetic fields indicates such fields could be used to detect and locate trapped miners. To be useful, the hardware of the system must meet a number of requirements, including small size, intrinsic safety, and rugged construction. Such hardware has been built, and the system has been tested by the Bureau of Mines and its contractors. These tests indicate that the electromagnetic method provides a practical means to locate miners in emergencies.

Jan 1, 1976

By Peter N. Blakey

The Kidd Creek orebody is a massive, base metal deposit with widths up to 600 feet and a strike length of 2,200 feet. The orebody dips eastwards at 70° to 80° and is amenable to sub-level open stoping. Five distinct zones (Fig.1) based on structural wall rocks and ore type parameters occur within the orebody. These zones are: Zone 1 - Mainly copper stringer ore Zone 2 - Massive pyrite ore Zone 3 - Massive sphalerite and chalcopyrite Zones 4 and 5 - Are banded and variable in nature Compressive strengths vary from zone to zone with an average of 22,000 p.s.i. and highs of 40,000 p.s.i. The mine is developed from a ramp extending from surface to the 1600 foot level and a 24 foot diameter, 3050 foot vertical shaft with main levels at 400 foot intervals. Sub-levels at 100 foot vertical intervals are driven from the ramp. transverse and longitudinal pillars to support the crown pillar left under the open pit. The layout allows an opportunity to enlarge or reduce pillar sizes and also to introduce fill on any horizon should conditions warrant. It also offers a large number of operating stopes spread through the various ore zones giving good ore blending characteristics. All ore from the stopes is handled through drawpoints using Wagner ST8 scooptrams with an average one way haul length of 650 feet.

Jan 1, 1976

By B. S. Ignasiak, Dennis Carson

"Laboratory study of the behaviour• of binary coal blends shows that coke strength is more closely dependent on the distribution of ""reactive"" oxygen in the parent coals than on their proximate, elemental or petrographic compositions, and it also appears possible to qualitatively predict coke strength from information about reactive oxygen. Proceeding from this recognition, experiments were carried out to test the possibility of preparing metallurgical cokes from binary blends in which non-caking hvb coal was substantially freed of reactive oxygen by prior heat t1•eatment under controlled conditions. The results indicate that such blends can yield surprisingly strong cokes and that coke strength may be amenable to further improvements by modifying the size distribution of the pre-heated component and the rate at which it is charred."

Jan 1, 1976

By Thomas A. Phillips

The Bureau of dines present, an economic evaluation of its process for the ferric chloride leaching of chalcopyrite concentrate. In this process, chalcopyrite concentrate is leached with ferric chloride solution to extract copper, which is precipitated from the solution by electrolysis while leaving sulfur in the residue. Spent liquor from electrolysis is treated with oxygen to regenerate the ferric chloride leach solution and the sulfur-laden residue is processed to recover the sulfur. A cost estimate was prepared for a plant capable of processing 600 tons of chalcopyrite concentrate per day, 330 days per year. The fixed capital cost, on an early 1.975 basis is $27,831,600 and the estimated operating cost is 9.9 cents, per pound of copper product. In comparison, the estimated operating cost of a smelter equipped with a citrate process plant to recover sulfur is about 13.6 cents per pound of copper product. No charge for chalcopyrite concentrate or credit for recovered byproducts has been included in either operating cost. The ferric chloride leach process, therefore, appears to be economically attractive.

Jan 1, 1976

By Leonard P. Vitt

This report details the work done to provide the U.S. Bureau of Mines with a customized Wireless Ground Monitor Wireless Ground Monitor was modified as follows: 1. A voltage regulator was added to prevent false trips with line voltage variations. 2. A Ground Wire Trap was designed to eliminate problems with parallel paths and provide spark suppression. 3. A three phase filter was designed to reduce 60 Hz leakage current. 4. A portable tester was designed to permit complete test of installed systems. 5. The design was modified to insure trip with an open ground wire in the event of a failure of any one semi-conductor component.

Jan 1, 1976

By V. F. Hollister, M. Bernstein

The Copaquire porphyry copper deposit lies in an island of pre-Upper Tertiary rocks surrounded by Quaternary and Upper Tertiary formations. The pre-Upper Tertiary appears to be a horst block in a typical basin-range setting. Major right lateral faults have been mapped within the range from Chuquicamata north to Copaquire. All of the major porphyry copper deposits known in the block (e.g., Quebrada Blanca, El Abra, Chuquicamata, and Copaquire) are on or are very close to one of these major right lateral shears. In the Copaquire district, an epizonal igneous complex with diorite, quartz diorite, quartz monzonite, and intrusive dacite (in that intrusive order) invade folded Jurassic shale. Mineralization appears spatially and temporally associated with the quartz monzonite porphyry and dacite. Alteration zoning shows two separate core phyllic (quartz-sericite-pyrite) zones within one large argillic (clay-pyrite) zone. A propylitic (chlorite) zone is peripheral to the argillic zone. No potassic zone has been recognized yet at the surface. The intrusive petrology together with the phyllic, argilic, and propylitic alteration zoning as well as the chalcopyrite-molybdenite ore mineralogy characterize this deposit as being typical of the Lowell and Guilbert model. Copper and molybdenum minerals appear to be most strongly developed in the phyllic zone, in either a breccia or a zone of veinlets and stockwork which trend dominantly northeast, conjugate to the district's largest fault, the Sulfato fault. The setting is basically similar to that found in some other porphyry copper deposits in this block, in that the ore body is a zone of veinlets which appear to be genetically related to movement on a major north-south trending right lateral fault. The presence of the major fault in each deposit, and the speculative hypothesis that paleo-subduction zone operated at the time of mineralization, suggest the possibility that mineralization took place when a major fault permitted escape to the surface of cal-calkalic magma generated in a subduction zone as proposed by Ericksen. A hypothesis such as that proposed by Ericksen explains the near linearity of porphyry copper deposits in Chile, since the source of calc-alkalic magma would exist at about the same distance from a paleo-trench throughout the length of the paleo-subduction zone.

Jan 1, 1976

By C. R. Larsen, R. G. Bradburn, A. L. Mular, B. C. Flintoff

"Following the acquisition of grinding- circuit data at Brenda Mines Limited for simulation purposes, various methods to mass balance the circuit were evaluated. Unfortunately, different answers were obtained for different balancing methods. A data-adjustment procedure was therefore developed. The procedure assumes that all ""best-fit"" criterion. A search is conducted with the minimum set of water masses and narrow size ranges of solids around the circuit (called search variables), which permit the computation of any of the measured data points. Search variables are used to calculate measured values for an objective function of the form: [ ] where W, is a weighting factor•, M1 is a measured value and M. is a corresponding . value calculated from search vanables. The best values of M, are those which minimize R. The above method is compared with others that have been utilized for mass balancing circuits."

Jan 1, 1976

By Ellis J. Brien, O&apos

This afternoon I would like to talk to you about coal, its defi- nition, its constituents, its associated minerals and present and future coal preparation techniques. This discussion will be an attempt at presenting sufficient coal preparation and associated fundamentals to enable you to understand "deep cleaning", a more advanced beneficiation process which will be commercially practiced in the near future.

Jan 1, 1976

By John R. Duda

The Bureau of Mines is developing estimates of capital investment, operating cost, and selling price for underground mines producing bituminous coal by various mining methods. This study estimates the required capital investment, operating costs, and selling prices for two hypothetical mines designed to produce 1.5 and 3 million tons per year (MM tpy)3 by using a longwall system in conjunction with a continuous-mining system. The coal properties being mined are assumed capable of sustaining a 20-year product Lon period. Wages and union welfare payments used in this study are those in effect as of December 6, 1975, as set forth under the National Bituminous Coal Wage Agreement of 1974. Costs of materials and equipment are based on January 1976 Indexes. Initial capital investment is $33.8 million for the 1.5-MM-tpy mine and $61 million for the 3-MM-tpy mine. Capital requirements, including deferred investments, were estimated to be $52.5 and $94.2 million, respectively. These figures translate into total capital investments of $34.87 and $31.31 per ton of annual production. Assuming a desired 15-percent discounted cash flow rate of return, with Federal income taxes at 50 percent, the required selling prices would be $12.06 and $10.94 per ton for the 1.5- and 3-MM-tpy mines.

Jan 1, 1976

By Gary M. Larwood

A study was conducted to determine the amount of transportation equipment required to move 1.2 billion tons of coal in 1985. The Federal Bureau of Mines has investigated a majority of the various practices associated with coal transportation and has gathered information on the regional origin-destination patterns, methods of movement, equipment stocks, rate structures, and operational capacities of the most important modes of coal haulage currently in use. Models employing this information were developed to establish the origin-destination pattern for 1985 coal shipments, estimate practical tonnage capacities for selected coal hauling rivers, develop modal shares for both unconstrained (by river capacity) and constrained cases, and present a range of coal transportation equipment requirements for given sets of constraints. Projected transportation shares for rail ranged from 63.7 to 72.3 percent, and for river, from 8.8 to 16.4 percent. Other coal transportation modes (truck, conveyor, Great Lakes, and tidewater) are expected to retain their historic shares of coal traffic. Equipment estimates were made for 1985 rail and water transportation based on 1973 average and best practice. Average practice estimates ranged from 675,000 to 710,000 hopper cars (100-ton) and 3,100 to 5,940 barges (1,400-ton equivalent). Best practice estimates ranged from 126,000 to 142,000 hopper cars and 1,750 to 3,450 barges. Full implementation of 1973 best practice by 1985 is unlikely.

Jan 1, 1976

By Alan K. Burton

Having estimated fleet size and productivity, the next and final step in this analysis is to introduce cost. Total cost of the system is a combination of capital, operating, and any other external costs that may be incurred. The last item, external costs, can be particularly important when comparing one size or type of truck fleet with another. Once costs are calculated, it is possible to compare systems on a total cost, cost per hour, cost per ton, or any other basis that is required. Methods used vary from simple owning and operating costs to sophisticated discounted cash flow analyses involving optional fleet replacements over the project life.

Jan 1, 1976

By Philip C. Good

Unweathered low-grade calcareous phosphate rock of about 20 percent P205 grade was beneficiated by a procedure in which contained carbonates were decomposed by calcination at about 1,000° C. On quenching the hot calcined rock in water, major impurities passed into a slime phase which was mechanically separated from a coarser concentrate fraction. Phosphate rock of two major types was beneficiated in batch-scale studies yielding concentrates containing 31.3 to 38.5 percent P205 at recoveries ranging from 54.8 to 86.0 percent. The samples were from two major deposits in India. Processing information gained was utilized in extending the work to a continuous-flow system, processing phosphate rock at rates up to 157 kilograms per hour. Superphosphate exceeding 16 percent water-soluble P205 was prepared from concentrate derived from continuous-flow processing.

Jan 1, 1976

By William R. Opie

Considerable effort by extractive metallurgical research organizations has been devoted recently to developing pollution-free processes for treating copper sulfide concentrates. A number of hydrometallurgical processes are in various stages of development and have been subjected to economic evaluation. Although the hydrometallurgical approaches eliminate SO2 emission problems, they create objectionable effluents and residues, make precious metals recovery complex, and are not very efficient for treating chalcopyrite. Furthermore, they are consumers of considerably more energy than pyrometallurgical processes which produce blister copper that can be electrorefined to commercial grade cathodes. This paper discusses a pyrometallurgical process under development by AMAX Base Metals Research & Development, Inc. (1) The original objectives set for the process development were: 1. Sulfur control of at least 95% with elimination of fugitive emissions. 2. Copper and precious metal recoveries as good as or better than conventional processes. 3. Low capital cost. 4. Competitive operating costs 5. Use of known proven technology where applicable.

Jan 1, 1976

By J. F. Oyler

One of the most important applications of material handling engineering technology occurs in iron ore pelletizing plants. Although the total installed value of material handling systems in such plants, as well as in many other modern process plants, is a relatively small percentage of the value of the plants (5 to 10% perhaps), their significance is considerable. Indeed this significance equals that of the more expensive process equipment in the plant, since the failure of any component of any system causes, shutdown of the entire plant. It is because of this increased consequence of failure that design criteria for process plants are more conservative than for conventional material handling applications. A second unique characteristic of material handling technology applied to process plants in general is the variety of materials being handled and the wide range of problems this variety presents. One extreme encountered is the stickiness of filter cake. Here the entire design effort is aimed at keeping the material flowing through transfer points and preventing it from building up on components. Another extreme is the conveying of fired pellets - hard spheres with a very flat angle of repose and abrasive characteristics severe enough to make conventional wear plate materials impractical.

Jan 1, 1976