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By M. E. Kravetz, E. B. Milosavljevic, L. R. Solujic

Gold ores containing high levels of cyanide-leachable copper minerals, such as covellite, chalcocite, bornite, malachite, azurite, and cuprite have traditionally been difficult to treat economically. A novel hydrometallurgical process for efficient treatment of cyanidation effluents obtained by leaching high copper gold ores (or concentrates) based on the use of Cherokee Chemical organo-sulfur reagents (OSR-A, OSR-B and OSR-C) is described. The process may be modified to selectively precipitate silver, or completely recover silver, copper and zinc, as well as regenerate significant amounts of cyanide. For example, data obtained demonstrate that 100 % of silver and zinc, and 98 % of copper were removed from the solution, while 80 % of cyanide was recovered. The treated barren solutions exhibit enhanced gold and silver leaching kinetics compared to the untreated controls.

Jan 1, 2004

By Philip R. Whitney

Precipitation of hydrous manganese/iron oxides in streams can be readily observed in much of the northeastern United States, and is apparently a common phenomenon in other glaciated areas (1,2,3,4). Deposition occurs when Mn-enriched ground or surface water enters the stream environment and encounters conditions of higher Eh and pH. The preferred location for deposition is in shallow, fast-flowing segments of stream; the oxides usually form as coatings on detrital material. The experimental studies of Hem(5) and a considerable amount of work on marine and lacustrine manganese nodules(6,7,8) suggest that given an appropriate substrate, deposition of manganese oxides may occur even when the water is undersaturated in Mn with respect to the oxides, due to catalysis by organisms or by silicate, carbonate, or ferric hydroxide surfaces. Once begun, manganese oxide deposition is autocatalytic (9). Studies of the distribution of oxides in various size fractions of stream sediments indicate that the thickness and prevalence of the coatings commonly increases with increasing particle size in the coarse sand and gravel size ranges. Figures 1 and 2 show a fairly typical distribution of leachable metals over nine size fractions from 62 microns up to 8-16 mm. The leaching solution used was neutral ammonium citrate with hydroxylamine hydrochloride, which effectively removes the oxide coatings without significantly attacking most silicates, carbonates or sulfides. Figure 1 shows metal content as ppm of the total sample; figure 2 shows' the same data as micrograms of metal per square millimeter of surface area.

Jan 1, 1975

By Philip R. Whitney

Deposition of manganese oxides is of common occurrence in streams of the north-eastern United States. Manganese entering the stream in solution in ground water is oxidized and deposited in the stream, primarily as oxide coatings on detrital materials. Prevalence and thickness of the coatings increases with increasing particle size in the coarse sand and gravel range. A number of metals of economic and environmental interest are enriched in these coatings. Samples of 8 to 16 mm pebbles were obtained from two streams in the Shawangunk region of southeastern New York State, an area of know71 Pb/Zn/Cu mineralization. The pebbles were leached to remove the oxide coatings and the resulting solutions analyzed for Mn, Fe, Pb, Zn, Cd, Cu, Co, and Ni. Trace-metal/Mn ratios in these streams are generally one to two orders of magnitude higher than regional background values. Profiles of trace- metal/Mn ratios along the streams show considerable detail; contributions of metals from tributaries are sharply reflected in the metal content of the oxides. The potential advantages and disadvantages of using coated stream gravel for geochemical prospecting and environmental studies are discussed in the light of these and other data.

Jan 1, 1976

By Waldemar Dyrssen

I CANNOT agree with the results obtained by the committee. The gas obtained per pound of coal is too high in calorific value and the amount of gas is too large. It should require about 190 cu. ft. of oxygen per 1000 cu. ft. of gas and not 75.14 cu. ft. as is given on page 35 of the report, or about 2.5 as much. Calculating the calorific value per pound of coal from the coal analysis given on page 34, according to Dulong's formula, the gross and net calorific heat are about 12,890 and 12,430 B.t.u., respectively, which is about 18,600 and 17,900 B.t.u., respectively, per pound of carbon in the coal. This is an unusually high figure, as ordinary bituminous coal has less than 18,000 gross B.t.u. per pound of carbon. Assuming, however, that the calorific values calculated are correct, the following results are obtained. On page 35, the net calorific heat in the gas per pound of coal is given as 12,856 B.t.u. and calculation for the gross value shows it to be about 13,950; in other words, the gas contains (13,950 - 12,890) ÷ 12,890 = 8.2 per cent. more heat than the coal on a gross basis and (12,856 - 12,430) ÷ 12,430 = 3.4 per cent. more on a net basis. The calculated amount of carbon in this gas per pound of coal gasified is about 0.688 lb., or nearly the same as in the coal (0.6972) and the gross and net calorific heat per pound of carbon in the gas are 20,300 and 18,710 B.t.u. which are 9.1 and 4.5 per cent. more than for the carbon in the coal. These calculations show that the results given in the report cannot be correct. There must always be some heat losses in the gasification of coal, in radiation from the producer, and in sensible heat in the gas; there must be some loss of carbon in ashes and in soot, and there are always tarry vapors in the gases that are condensed and do not appear in the fixed gases. From the excessive heat in gas per pound of carbon contained therein, it can also be concluded that the percentage of hydrogen in the gas is too high, from which it follows that the moisture decomposed is too high and that more free oxygen must be used.

Jan 11, 1924

By P. Hanniala, S. Harkki, B. Andersson

"The Outokumpu Flash Smelting Process was put into commercial operation in 1949. The original process was based on the use of preheated air. The adoption of oxygen enrichment of the flash smelting process air in 1971 was a crucial improvement, especially for smelter capacity. Oxygen use according to the Outokumpu method means selection of the most suitable oxygen content for the process air on the basis of the concentrate composition, capacity and matte grade. The furnace is not bound to only one fixed degree of process air oxygen enrichment; the oxygen content may vary from 21% to 100%.The degree of oxygen enrichment and the matte grade are the most important process parameters in reducing the operating costs of the whole system, including smelting, converting and acid production. Raising the matte grade allows more efficient use of the reaction heat in the smelting stage. Sulphur is recovered in a high-strength, constant-flow gas stream. The blowing time in the converters is reduced considerably.Oxygen enrichment and high matte grade improve the energy efficiency of the plant complex as long as the surplus heat from the system is not wasted. When the operating costs and investment expenditure of the total plant complex are taken into account, a reasonably high degree of oxygen enrichment (35-80°/0) is the best alternative in practice. Existing operating smelters have achieved single-stream thoughputs of over one million t/yr of concentrates.HistoryLocal capital and operating costs as well as lower energy consumption per unit of metal produced are now the crucial factors in determining the choice of process. In the late I940s, the same reasons prompted Outokumpu to develop the original flash smelting process in Finland. Between 1900 and 1940, very few technical changes in pyrometallurgical copper production took place. However, the commercial use of the flash smelting technique began a new era in metal extraction from sulphide concentrates. The characteristic feature of the method is that the concentrate is smelted in suspension: this offers the most favourable conditions for heat and mass transfer, and thus high reaction velocities. Outokumpu was the first to exploit the known advantages of flash smelting in a commercial installation at Harjavalta in 1949."

Jan 1, 1982

THERE was presented for discussion at the February (1924) meeting of the American Institute of Mining and Metallurgical Engineers a report of a committee named by the United States Bureau of Mines on the Use of Oxygen or Oxygenated Air in Metallurgy and Allied Processes. An extensive discussion followed the presentation of the papers, which dealt with the matter entirely from a theoretical standpoint. The results of only a limited, amount of research work were presented and discussion centered around the question: "Is it worth while to carry on research and expend money in determining what can be done in metallurgical operations if cheap oxygen can be produced?" More experimental work on this subject has not been done because metallurgists have not thought that cheap oxygen could be produced, thinking "Why waste time experimenting. with a reagent that cannot be procurable in commercial quantities?" While the producers of oxygen have said " Why waste money on the development of a process to make cheap oxygen unless we know that there will be a market for it when produced?" The discussion brought out two facts: First, many metallurgists believe that their operations can be conducted more economically if they can add oxygen to the air used in various smelting operations; second, a number of chemists believe that relatively pure oxygen can be produced in metallurgical quantities at very small cost. Negative views were expressed to both propositions, but so overwhelming was the view that if cheap oxygen could be produced, many metallurgical operations would benefit that it was voted to recommend that the Bureau of Mines continue its research work, pending the production of the cheap oxygen hoped for. Few experimental data being available, and the attitude of disputants being based largely on theoretical considerations or on reactions and temperatures incapable of positive proof, it has been decided to publish only a summary of the papers and discussions arising from their presentations. It is believed, however, that this will be sufficient to show the lines of reasoning that have been adopted, and be a guide to others who are to make their own calculations or engage in experimental work.

Jan 11, 1924

R. H. SWEETSER, Columbus, Ohio.-My experience with oxygen in the blast is limited to the use of one tank of oxygen in melting out a chilled anthracite furnace. The oxygen was introduced through the peephole of the tuyeres and opened them in a very short time. We could see that an intense heat was generated just as soon as the oxygen struck the fuel. I agree with the committee in respect to many. of the things that will probably happen in the iron blast furnace, particularly as to the speeding up of the operation. There will undoubtedly be changes in the shape of the blast furnace, and if I were designing one today I should hesitate about the form to give it. Mr. Kirby suggested that the proper shape might be rectangular. Some of the men of the American Rolling Mill Co. have reached the same conclusion, as a result of Royster and Joseph finding that the combustion zone in the front of the tuyeres extended into the furnace less than 40 in. If this is correct, we might build the furnace with an elongated rectangu-ular hearth having the shorter dimension a little over 80 in. and secure. a very intense temperature in the hearth. The committee's report said, "With more oxygen available in the blast, the smelting ability of the furnace will be increased beyond the' speed of the preparing zone in the shaft." This can be controlled by providing ample facilities for getting stock into the furnace, and this will be one of the biggest changes in blast-furnace construction. The question has been raised as to finding suitable refractories to withstand the increase of hearth temperature. If the furnace can be kept full, the descending column of cold material will probably take care of the increased heat from more rapid combustion of the coke. At the Cleveland meeting, in 1912, the iron blast-furnace men listened with great interest to a paper on copper slags; from Mr. Kirby's paper, it looks as though they might again have to go to the copper men, for many of the problems of the iron blast furnace of the. future that uses oxygen may be worked out on lines of the copper blast furnace.

Jan 11, 1924

By Arthur McKee

As HEAT is the controlling factor in all smelting operations, it will be .most helpful to make a new set of calculations on the basis of a .unit of fuel burned at the tuyeres to carbon monoxide. This ties together the only two factors in this operation that seem to be constant with respect to each other-the fuel at the tuyere and the air or its oxygen content. If our usual smelting operations, using natural air heated to reasonably obtainable temperatures, furnish sufficient heat above this critical temperature for the purpose, it is doubtful whether additions of oxygen will be of marked benefit. A unit of fuel will not produce a single heat unit more if burned with pure oxygen than if burned with natural air and less heat will be available per unit of fuel by the amount of heat that is carried in by the preheated nitrogen in the natural air of the blast. This is more readily apparent if considered in connection with a heat balance sheet per unit of fuel at the tuyere. The heat available in the stack, as shown by the committee's report, is less by reason of the lower fuel consumption per unit of output assumed in this calculation. The balance sheet for enriched air shows a shortage of the amount of heat usually carried in with the blast; also a lack in the hearth of sufficient heat to melt the 18 per cent. additional metal and slag and in the shaft to the extent of the shaft heat required to heat the 18 per cent. additional stock to the melting point. These heat deficiencies are provided for by the largely increased conversion of carbon monoxide into carbon dioxide, which is assumed, as shown by the following figures: Calculation for natural air: Carbon dioxide 18.4 per cent. (after having eliminated the carbon dioxide from limestone) and 20.8 per cent. of carbon monoxide. Calculation for oxygen-enriched air: Carbon dioxide 28.4 per -cent. and carbon monoxide 23.2 per cent. (no carbon dioxide from stone as lime is charged instead).

Jan 11, 1924

By Charles Hart

THE-report of the advisory committee to the U. S. Bureau of Mines, on the use of oxygen in metallurgy, brings to the art of steelmaking a radical change in the method of operation of the many processes of this widely diversified industry. When completed, the improvements suggested by the use of oxygen will be far-reaching and revolutionary. As to its effect upon the iron blast furnace, the use of oxygen should be as important as any of the improvements in the art of iron smelting and should take its place beside the German "Hochofen" which was the first attempt to use a shaft for preheating the stock, and beside the hot-blast stove which heated the blast before entering the furnace. The former initially produced molten iron and the latter made unbelievable economies in fuel consumption. It would seem that a discussion of this paper would be best served by following the conclusions on page 17 of the report and analyzing them seriatim. 1. (a) A decrease in costs, due to higher smelting heats is only attainable by taking full advantage of these heat units. This is affected by, adding additional ore to the burden which in turn increases production and yields a larger tonnage upon which to figure costs. These economies appear in decreased cost per ton of iron for coke and limestone consumed, and a proportionate reduction in all items entering the cost above materials. We can take no exception to these. premises. It is a well-known fact that silicon in the iron and sulfur removed" are the result of hearth temperature, other matters remaining constant. To a practical furnaceman the most satisfactory manner of controlling a furnace is by increasing the burden. It is not necessary, therefore, to consider the question of refractories, for this discussion presupposes that the temperatures in the furnaces will be kept practically where they now stand for making the various grades of iron. The same situation obtains in regard to the height of the furnace, which, as intimated, was one of the great strides forward in this art, and its abandonment at this time, is inconceivable. The shaft prepares the stock so that at each succeeding step we have exactly the proper conditions. All the stock entering the furnace is dried and preheated. The limestone is calcined by sensible heat practically at no cost, for this is sensible heat which otherwise would pass from the top with the gases. Under normal conditions, external calcination has been attempted without tangible results. Under normal average furnace operation, limestone requires about 760 B.t.u. for calcination and the loss of sensible

Jan 11, 1924

By I. Iwasaki

Differential flotation of bulk sulfide concentrates from Duluth gabbro into copper and nickel concentrates could be implemented by first destroying the hydrophobic coating with ozone and then by recovering the copper concentrate with a copper-selective collector. A number of parameters involved in the use of ozone; such as the sequence and the conditioning time of ozone treatment, the pulp pH, and the level of xanthate addition, were investigated.

Jan 1, 1985

By J Shillabeer

Paste backfill has been adopted by many underground mines in North America and a growing number elsewhere. Paste offers a number of advantages. This paper suggests wider applications of paste technology to surface tailings disposal and water management. Advantages include improved overall stability, potentially easier permitting, smaller water inventory and earlier reclamation. It may be feasible to use the same plant to prepare backfill and to dispose of surplus tailings. Example cost models are given to illustrate potential applications.

Jan 1, 1998

By Shillabeer J

Paste backfill has been adopted by many underground mines in North America and a growing number elsewhere. Paste offers a number of advantages. This paper suggests wider applications of paste technology to surface tailings disposal and water management. Advantages include improved overall stability, potentially easier permitting, smaller water inventory and earlier reclamation. It may be feasible to use the same plant to prepare backfill and to dispose of surplus tailings. Example cost models are given to illustrate potential applications.

Jan 1, 1998

By T. Malterer, R. Adams

Peat has unique physical, chemical and biological properties. These include high porosity and water retention, high absorption and adsorption properties, and high buffering capacity. In addition, peat is a rich medium for microorganism activity. Peat has traditionally been used for horticultural purposes. However, in recent years, peat is increasingly being used for waste- treatment purposes. This paper reviews North American literature on the use of peat for waste-treatment applications. The literature indicated that peat has potential for use in the treatment of municipal and on-site effluents, for the removal of low concentrations of heavy metals and for the adsorption of liquid and gaseous hydrocarbon pollutants.

Jan 1, 1997

By M. E. Holuszko

Although there is no widely accepted rigorous definition of handleability, the handling coal characteristics often referred to as handleability define whether a coal has the ability to flow unhindered through the processing and transportation systems. The handleability may be severely affected if fine coal particles tend to aggregate. In the pelletization process, the rolling action of the drum is applied to bring the individual particles into proximity with each other so that they can aggregate and form pellets. Because of apparent similarities between these two processes, the pelletization tests are carried out in parallel to the handleability tests in this project, and the pelletization results are used to explain coal handleability properties.

Jan 1, 2004

By D. C. Martin

Stability analysis and design of open pit slopes is generally based on an assessment of rock cores and rock exposures on surface or in underground workings. Until recently, core for such assessments has been readily available from exploration drilling programs. In recent years, many exploration programs have reduced the amount of coring and rely more heavily on less expensive rotary or percussion drilling programs where drill cut-tings are sampled for geological evaluation and assay. This paper describes a technique whereby costs of the geotechnical investigation are reduced by using rotary percussion drilling information to assess geotechnical properties of the rock mass. A limited number of core holes are required to establish basic relationships between the mechanical properties of core and the characteristics of the percussion chips. These properties are used as a basis to extrapolate the coring information to encompass the entire area of the rock mass for which percussion drilling information is available.

Jan 1, 1990

By H. McIntire, R. Preston, D. Roberts

"Repeated 3D terrestrial digital photogrammetric surveys were conducted on hard rock mine pillars to characterize damage during secondary mining. Damage was quantified by measuring Depth of Failure, DOF, from spalling material and crack intensity, D21. These damage values were then correlated to visually assessed Pillar Damage Ratings and numerically estimated pillar loads. Measured damage, particularly Depth of Failure, correlated well with Pillar Damage Ratings. Numerical predicted pillar stress vs. damage exhibited the expected trend with shorter pillars being stronger. High resolution photographs and 3D models were also used to characterize damage mechanisms, which were found to be structurally constrained. INTRODUCTIONThe Doe Run companies Southeast Missouri Operations consists of six room and pillar lead mines (Brushy Creek, No. 29, Casteel, Buick, Fletcher and, Sweetwater) (The Doe Run Company, 2014). The mines are hosted in the Carbonates and Shales of the Viburnum trend with mining depths varying from 180 to 300 m. Pillars are commonly cut square with widths varying from 7 to 9 m and heights generally between 5 and 18 m, resulting in extraction ratios between 80 and 85% (Roberts et al., 1999). Figure 1 presents an idealized stratigraphic section through the Viburnum trend and Figure 2 shows a plan view of the five areas spanning the three mines included in this study."

Jan 1, 2015

By Gross A. E, Neville P. C

Over the last decade heap leaching with a cement binder has allowed the economical leaching of low grade gold ores of less than 2 grams/tonne. Despite the continued use of cement as an agglomeration aid, numerous leaching operations continue to experience channelling and blinding problems. This paper describes the development of polymeric agglomeration aids, Extract Ore 9760 and 9960, which when used as a partial cement or lime replacement can result in a higher rate of and/or a higher total gold recovery in plant evaluations. Extract Ore in combination with cement acts to produce a stronger agglomerate which results in reduced fines migration during leaching and helps prevent blinding of the heap. Both of these changes will give a more even flow of cyanide solution throughout the heap and improve solution contact efficiency within the ore. Polymeric agglomeration aid programs are now being used by heap leaching operations in three different countries. A summary of benefits identified at some of these operations are presented in this paper.

Jan 1, 1994

By Matthias Pulsfort, Markus Herten, Rolf Breitenbücher, Claudia Fierenkothen

"Secant bored pile walls are frequently used as retaining walls in watertight excavations to cut off ground water aquifers and are subsequently exposed to ground water pressure loading in addition to earth pressure loads. To avoid extra loading from ground water pressure in soils with limited permeability, it can be useful to arrange drainage windows within secant bored pile walls by means of single primary piles filled with porous concrete without reinforcement cages, which usually have to be placed within casings in submerged conditions. This concrete may serve as a horizontal arch between the reinforced secondary piles, but it must however enable vertical draining between load carrying construction beams and the bottom of the excavated pit. In the present submission, the results of several material tests with different mix designs of porous concrete with certain admixtures are presented together with first and important experiences with a wellsuited placement technique for porous concrete casings in small- and large-scale laboratory tests. It can be shown that the usual placement technique of employing a tremie pipe is not suited for this type of unsaturated concrete. Instead, pouring with a particular kind of cone and an open conductor pipe has proven successful.Additionally, special testing methods are currently developed for judging the suitability of the desired concrete mix design in preparation for an upcoming real-scale experimental pile installation in Germany.1. INTRODUCTIONRetaining walls are often built as secant bored pile walls in order to cut off aquifers saturated with groundwater against deep building pits. In soil types with relatively low permeability, however, seepage flow may be permitted, because the quantity of groundwater to be pumped is limited, and the range of influence due to dewatering is not reaching far. Thus, the hydrostatic water pressure on the retaining wall can be considerably reduced, leading to a more economic design (see Fig. 1).In unsaturated soils, the dewatering effect can be achieved by means of porous pervious concrete with a uniform grainsize distribution of aggregates typically placed in a borehole under dry conditions. With regard to the use of porous concrete to be placed at the ground surface in road pavement applications, experience is sufficiently available as shown, for instance, by Neithalath et el. (2010). However placing the porous concrete in a bored pile using a tremie pipe under submerged conditions is a challenge and a completely new approach, because there is a considerable risk of segregation of the fresh concrete, separating the cement lime paste from the aggregates when in contact with water and leaving only neat gravel on the one hand as well as a tight cemented area at the bottom of the pile on the other hand. If this concrete placement is successful, additional dewatering by means of extra wells behind the retaining wall can be avoided. Of course the seepage flow in the adjacent soil needs to be regarded seriously."

Jan 1, 2017

By Jacob A. Jordan, Rachel M. Garrett, W. Robert Thompson III, James C. Pegues

"Environmental control structures for an existing fossil-fuel power generating plant near Pascagoula, Mississippi faced several design and construction challenges due to space limitations. Conceptual design calculations indicated drilled shaft foundations would likely be needed to accommodate the higher lateral and compressive loads expected under some structures. A maximum drilled shaft diameter of 60-in was preferred due to various space constraints. Smaller diameter drilled shafts were deemed acceptable under structures with more moderate loading conditions. Furthermore, augercast piles were considered where lateral loading was not a governing design criterion.A series of SPT borings and CPT soundings revealed a subsurface profile consisting of soft to firm clays to depths of about 25 feet, underlain by fine to coarse sands. The apparent relative density of the sands was loose just below the clay, with increasing density noted with an increase in depth.Analysis of conventional drilled shafts indicated needed capacities could not be achieved unless shaft lengths exceeded 100 feet. Post-grouted drilled shafts were therefore evaluated to decrease shaft lengths by taking advantage of increased base capacity. Analyses showed the higher predicted capacities needed for some structures could be achieved with this approach. Final design for the project incorporated 90-ft long, 60-in diameter post-grouted drilled shafts; 60-ft long, 48-in diameter post-grouted drilled shafts; and 60-ft long, 48-in diameter conventional shafts. Both post-grouted and conventional trial shafts were installed as part of the detailed design phase of the project. Each trial shaft included CSL integrity testing and Osterberg-cell bi-directional load testing. Load test results revealed that additional base capacity was, in fact, achieved through the post grouting process. The 60-in diameter post-grouted production shafts and the 48-in diameter conventional production shafts have been installed at the site, with the project that would utilize the 48-in diameter post-grouted shafts deferred at this time."

Jan 1, 2015

By A. C. Maffei, M. V. Lins Gonçalves

Ore concentration processes are a function of the respective particle size distribution; however, most of them do not use an efficient operational method to determine it. From a practical point of view, the surface area can be calculated for any size analysis. The physical principles for determining the size distribution in each method are different, for the results will not be compared directly. During the development of the research project of the sulfide ore flotation method using inert gas, different methods for calculating the specific surface area have been used, such as sieving, sedimentation/X-ray, laser diffraction and BET, i.e. specific techniques to determine the specific surface area of copper and gold ore concentrates. The statistical method of minimum squares was used to achieve practical correlation equations among them. Regarding these two ores, if one knows a specific surface area value using one of the methods above, one can predict the result of the others. Diminution of the reagent consumption was so controlled and it could be even decreased by using inert gas.

Jan 1, 2003