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By Elmer R. Kaiser

COMBUSTION of coal and transfer of heat from flames and gases to boiler surfaces continue to be of great interest to engineers here and abroad. Numerous investigations have been in progress to improve furnace and boiler performance and economy. The importance of better understanding of the processes and opportunities for improvement is apparent when it is remembered that heat from at least 500 million tons of coal a year the world over is being transferred to boiler water at efficiencies ranging mostly between 50 and 90 pct. Even slight gains in efficiency, economy, and labor saving become very significant when multiplied by the enormous quantity of fuel consumed. Also the competitive position of the large coal, oil, and gas industries in satisfying the fuel consumers is greatly affected by the achievements made through technical progress with each fuel. This paper is part of a continuing activity of Bituminous Coal Research, Inc., to extend the knowledge of coal utilization for steam generation and to seek promising directions for future research and development in cooperation with others. Particularly in the latter regard, numerous interviews were held during the last three years to seek the experience and advice of boiler and combustion-equipment manufacturers, electric-utility executives, and fuel engineers. A wealth of published information was also reviewed, which together with the interviews pointed to the advisability of further work on ash and sulphur control. For the present purpose a number of factors important to efficient heat liberation and recovery have been grouped as follows: 1—combustion, temperatures, and rates of heat liberation; 2—radiation, convection, and furnace and boiler configuration; 3—sponge ash, slag, and hard-bonded deposits; 4— low-temperature deposits and corrosion (cooling flue gas below dew point and air-pollution control); 5—the limitations of coal cleaning and boiler size and cost as related to fuel characteristics; 6—future possibilities and conclusions. The development of combustion apparatus for power boilers is progressing at a lively pace. There has been no letup in improvements in design of pulverized-coal-fired boilers, and there is a strong trend at present toward improving dry-bottom units. Spreader stokers with overfire jets and dust collectors as standard equipment are gaining favor. Less than 10 years in commercial use, cyclone burners are going into numerous installations here' and abroad.' Underfeed and traveling-grate stokers have long since been developed for heavy-duty operation, yet new developments in overfire jets and humidification of air blast have improved their performance. A water-cooled vibrating-grate stoker of German origin is being introduced into the United States and Canada." The primary objectives of an ideal coal combustion device are: capacity to burn the variety and sizes of coals likely to be economically available during the life of the unit; capacity to burn the coals automatically for a wide load range and rapid load fluctuations and to burn the coals completely to CO2, H2O, and SO2, which means without smoke and cinders, or carbon in the refuse; capacity to control and discharge all the ash in final granular form without ash adhesion to walls or tubes, and without flue dust; minimum furnace volume; minimum labor and maintenance; low initial and operating cost. Regardless of the method of burning, the gaseous products of coal combustion are N2, CO2, O2, H20, and SO?. By way of illustration, the coal analyses in Table I is assumed from an installation described by E. McCarthy.' When coal is burned with 20 pct excess air (theoretical air, 9.23 lb per lb of coal), the quantities of combustion gas shown in Table II are produced. In addition, the gases carry particles of fly ash, unconsumed cinders, soot particles, and small but significant amounts of vaporized oxides and sulphates of sodium, potassium, lithium, phosghorous, iron, and other metals. In recent years, germanium, one of the rare metals found in coal, has been shown to oxidize and vaporize at combustion temperatures and to be concentrated by reconden-sation at lower temperatures." Pulverized coal and cyclone flames" have peak temperatures of 3000' to 3500°F. Temperatures in fuel beds of large underfeed stokers reach maxima of 3000°F, sufficient to fuse almost any ash and to volatilize some of it. These peak temperatures are above the optimum necessary for rapid combustion, but they hasten heat transfer for ignition as well as boiler heat absorption. Furnace and gas temperatures increase with combustion air preheat. Low excess air has the same effect. Fine coal pulverization and highly turbulent combustion shorten the distance for fuel burnout, increase flame temperature, and speed up heat transfer. Rates of combustion of pulverized coal exceeding 200,000 Btu per cu ft per hr have been demonstrated in atmospheric gas-turbine combusters,

Jan 1, 1955

By M. M. Wong, D. E. Couch, G. M. Martinez

The Bureau of Mines electrowon hafnium metal with an average oxygen content of' 150 ppm at 700°C from an electrolyte containing 27 wt pct LiCl, 62 wt pct RbCl, and 11 wt pct HfC14. The average anode and cathode current efficiencies were 90 pct at anode and initial cathode current densities of 86 amp per sq ft. Haf-nium metal with an average oxygen content of 440 ppm was electrowon at 800oC from an electrolyte containing 90 wt pct KC1 and 10 wt pct HfCl4. The average anode and cathode current efficiencies were similar to those obtained in the LiCL-RbCl-HfCl, electrolyte. The chlorine gas given off at the graphite anode was vented through either a silica or a graphite tube to prevent cell corrosion. THE current method for the commercial production of high-purity hafnium is the thermal decomposition of Hfl4.1 The iodide method is not adaptable to continuous process techniques. Nettle, Hiegel, and Baker2 studied the electrorefining of hafnium from hafnium sponge containing 800 ppm oxygen. They failed to obtain hafnium with 600 ppm oxygen in their initial deposits and obtained AEC specification for oxygen only after 75 pct of the soluble hafnium had been removed from the electrolyte. Calculations using their data indicated this was approximately 4 lb of hafnium. The electrolyte was then used to produce approximately 3 lb of hafnium with a low oxygen content. However, no data are shown concerning the amount of anode material initially used or what percent of it was dissolved, therefore, results are not suitable for evaluation of a continuous operation. In general, it was not possible to consistently obtain low oxygen content metal with the electrolytes described by Nettle, Hiegel, and Baker. Wong, Hiegel, and Martinez3 investigated the electrorefining process for hafnium and showed that even by strict control of electrolyte composition only relatively low oxygen reduction could be obtained. The oxygen contained in the hafnium anode material tended to transfer to the cathode deposit and only a limited purification was possible. Both the "iodide" and the "electrorefining" processes depend upon hafnium sponge as a starting material. The sponge is normally produced by magnesium reduction of HfC14 ' and does not meet AEC specifications for hafnium metal. Since only 30 pct of the anode feed could be utilized3 in the electrorefining cells, the Bureau of Mines developed an electrowinning process. HfC14 was used as the feed material for the electro-winning process described in this report. Many of the electrolytes used in the electrorefining studies3 ap- peared to be suitable carrier-electrolytes for HfC14. However, in the initial studies on electrowinning, it was desirable to use electrolytes that had low solidus temperatures and could be operated over a wide temperature range to investigate parameters of the process. Therefore, electrolytes containing LiC1, NaC1, KC1, RbC1, CsC1, and HfC14, in various combinations were explored. EQUIPMENT Chlorinator. Hafnium carbide was chlorinated to produce HfC14 in the batch-type chlorination shown in Fig. 1. Chlorination temperatures were measured with a thermocouple placed in the center of the HfC charge. A flow meter was used to monitor the helium and chlorine. The exhaust side of the silica chlorina-tor tube was equipped with a flask for collecting organic material released during the initial heating of the HfC. The temperature of an internal heater, which extended from the HfC14 condensing flask to the hot end of the chlorinator, was adjusted to prevent the HfC14 from condensing before entering the collection flask. Helium and excess chlorine were exhausted through the lid of the collection flask to an aqueous NaOH solution. Sublimer. Initial studies were conducted using a sublimer, Fig. 2, made by placing a 13/8-in. OD nickel thimble 11 in. long, inside a 11/2-in. ID nickel bell 12 in. long, and locking it in place. This unit was loaded with HfC14 and partially immersed in the molten electrolyte for sublimation directly into the electrolyte. In another sublimer shown in Fig. 3, the HfC14 was contained in a "resin reaction flask". Quartz wool, previously heated to 600aC, secured between two nickel wire screens, was placed just above the HfC14 powder. The lid contained a vacuum outlet, a gage, an argon inlet, and an air-cooled pipe for condensing the HfC14. This sublimer was evacuated and heated. The sublimation temperature was not critical and the sublimer operated satisfactorily at all temperatures between 250" and 350°C. Electrolytic Cell. The electrolyte chamber, Fig. 4, was made of mild steel 8-in. schedule 20 pipe, 30 in. long. The exterior was metallized with a Ni-Cr alloy. The electrolytes were contained in a 16 gage nickel or iron liner with a nickel heat shield on top. The cell was heated by a resistance furnace. A 21/2-in. ID by 25 in. long air lock was connected to one port of a two-port cell cover assembly through a slide valve. The cover assembly of the air lock was electrically insulated from the cell and was equipped with a rubber sleeve that provided for the passage of the cathode lead. This allowed the cathode deposits to be removed and a new nickel cathode to be introduced without allowing air to enter the cell. A tube-rod assembly was bolted to the other port on the cell cover assembly and was sealed by a packing seal. The tube-rod assembly consists of a graphite

Jan 1, 1970

By R. V. Norris

It was stated in a former paper1 that an intensive study of anthracite costs was being made by the engineers of the United States Fuel Administration. The results of this study are now available and are offered as a supplement to the former paper. Anthracite costs as reported for the 6-months period, December, 1917, to May, 1918, inclusive, as compiled by the Federal Trade Commission, were used as a base, and charts similar to those described in the former paper were made for the standard white ash anthracite and for red ash and Lykens valley coals. Adjustments The adjustments of cost from a reported to a price-fixing basis, as described for the bituminous methods, were applied but showed only minor adjustments as necessary. The great spread in anthracite prices on the varyingsizes, which for the 6-mo. period under review ranged in average from $5.244 for nut to $2.074 for barley coal, makes the question of the percentage of sizes produced at the different collieries a vital one. The realization with the same prices for each size must be within very wide limits, when it is considered that the percentage of prepared coal reported from different collieries varied from over 80 per cent. to below 30 per cent. for fresh-mined coal. Hence, as the spread in prices for the various sizes must be predicated on some percentage, it is essential to find some method of adjustment to allow for this variation. The logical method of adjustment is to calculate actual costs to costs as of the standard percentage of sizes, so that the margin between the adjusted costs and the average realization shall be the actual margin for each colliery between its actual costs and actual realization due to its particular percentage of sizes. As a base for realization the actual percentage of sizes for fresh-mined coal for the 6-mo. period was adopted. This percentage is given in Table 1.

Jan 1, 1920

By C. O. Dale, W. J. Rundle

THE upper Mississippi Valley zinc-lead area was the first major lead-producing section in the United States. The lead ore, found near the surface in crevices, was relatively pure galena that could be smelted directly into lead, at first in log hearth furnaces and later in more efficient blast type furnaces. French Canadian fur traders encouraged the Indians to mine the lead ore and showed them how to smelt it into lead that had a high value for bullets1 Nicholas Perrot found lead ore on the Mississippi River bluffs near the junction of Wisconsin and Illinois and in 1690 established a trading post on the Wisconsin side of the river opposite the present site of Dubuque, Iowa.2 Shortly after 1720 discovery of Mine La Mott in Missouri diverted considerable attention from the Upper Mississippi area. Mining continued on a desultory basis with operations concentrated in the Galena, Illinois-Dubuque, area. In 1740 at least 20 miners were at work in the Fever River area around Galena and are reported to have shipped 2500 70-lb pigs of lead to Kaskaskia in 1741." Julien Dubuque established a mining and smelting operation in 1790 near the city that bears his name and was granted sole right to exploit the mining operations on the lands of the Sauk and the Fox Indians. He is reported to have produced 30,000 70-lb pigs of lead in 1805. Following the death of Dubuque in 1810 the Indians refused to let the white miners enter their lands, and little was done on the Iowa side of the river until the Indians were removed by treaty with the United States government in 1832." Early mining was entirely for lead but as the crevices were followed down, increasing percentages of zinc sulphide and zinc carbonate were encountered and at first discarded. Later a market became available for the zinc ores, and hand jigging devices were made to separate the lead, the zinc, and the rock or waste materials. The first record of zinc production from the area is for 1860. Production of zinc passed that of lead before 1900, reached a peak of 64,000 short tons in 1917, fell off rapidly and continually to about 2000 short tons in 1938, and since 1940 has ranged from 11,000 to 19,000 short tons. Lead has been of considerably less importance since 1900, and at present only about 10 pct as much lead as zinc is produced. Practically all of the zinc ore has come from orebodies that are rather flat and wide with considerable length as compared to width. Most of the early lead came from the crevice type deposit, but present production is from the predominately flat zinc orebodies. The Graham-Snyder orebody, scene of Eagle Picher operations, is practically all zinc with little or no lead being recovered. Marcasite, present in varying amounts, makes production of finished concentrates by gravity separation impractical. Satisfactory lead and zinc concentrates have been produced since flotation was introduced in the area in 1927. An acid recovery plant was operated for about 20 years after World War I, but it has been dismantled, and no recovery of the iron sulphides in the ores of the district is being made at the present time. In June 1950 there were three companies operating mines and mills, Tri-State Zinc Co., Calumet & Hecla Consolidated Copper Co., and Eagle Picher Mining and Smelting Co. The Vinegar Hill Zinc Co. had completed a shaft at a new orebody and had started to develop the mine which will supply the Cuba City mill. The Cuba Mining Co. was holding the Andrews Mine inactive. The Dodgeville Mining Co. was not operating but was exploring for additional reserves. Several small mines were selling ore to the Eagle Picher mill. A general area map is given in Fig. 1. The Eagle Picher Mining and Smelting Co. entered the area in 1946 with an active exploration campaign. Leases on a block basis were secured for the area south from the Wisconsin-Illinois line near

Jan 1, 1953

By J. D. Testerman

A statistical technique to identify and describe naturally occurring zones in a reservoir and to correlate these zones from well to well is described. The technique is particularly useful in describing a reservoir where cross flow between adjacent strata is important in determining reservoir behavior. Although it has been used primarily for permeability zonation, the technique is general and can be used to correlate any reservoir property or related data, such as the information contained in well logs. INTRODUCTION One of the first problems encountered by the reservoir engineer in predicting or interpreting fluid displacement behavior during secondary recovery processes is that of organizing and using the large amount of data available from core analysis. Permeabilities pose particular problems in organization because they usually vary by more than an order of magnitude between different strata. Due to the sheer volume, it is almost always necessary to group data and to use an average value to represent a number of measurements. Perhaps the most common method now used to group permeability data is the capacity-fraction technique, which ranks permeabilities in order of magnitude, regardless of the physical location of the permeabilities within the reservoir. The cumulative per cent capacity is plotted against cumulative per cent thickness. This plot is divided into an arbitrary number of zones, generally of equal thickness. Five zones (or averaged groups of data) usually are obtained, each of which is treated as homogeneous in subsequent calculations. The division so obtained has no physical meaning; strata in the same zone, calculation-wise, are usually not adjacent in the reservoir. Reservoir engineering techniques being developed will handle crossflow that occurs between adjacent communicating reservoir strata because of imbibition and gravity segregation. Since crossflow occurs between physically adjacent layers within the reservoir, a new zonation technique recognizing the actual location of strata within the reservoir is necessary. Similarly, the recognition of natural zones is important for predictions of oil recovery by processes involving diffusion. One such process is miscible displacement, where predictions of lateral diffusion within the reservoir must recognize the actual location of the invaded zones in relation to the rest of the formation. Natural zones must also be adequately recognized to account for heat transfer within the reservoir during thermal exploitation. Because of the complexity of the problem, statistics appear to offer the only practical hope of dividing a reservoir into physically-meaningful natural zones. This paper presents a statistical technique for identifying these natural zones and for ascertaining which ones are likely to be continuous between adjacent wells. The zones defined have minimum variation of permeability internally and a maximum variation between zones. The technique is general and can thus be applied to reservoir properties other than permeability. The method will guide the reservoir engineer in estimating which zones are likely to be continuous between wells. However, a statistical correlation based on permeabilities in two different wells is no guarantee that the zones so defined are, in fact, continuous. Rather, the assumption of continuity must be consistent with geological data concerning the depositional environment, as well as justified on the basis of engineering judgment in combination with statistics, just as judgment is required with conventional zonation methods. CALCULATION PROCEDURE The reservoir zonation technique is a two-step operation. The steps are individually described, and a sample calculation is presented in the Appendix. ZONATION OF INDIVIDUAL WELLS First, the set of permeability data at a single well is zone, into Zones. These zones are selected so that variation is minimized within the zones and maximized between the zones. The equations4,6 used to zone the data are where B = the variance between zones, , = the number of zones, i = the summation index for the number of zones, j = the Sumation index for the number of data within the zone, mi = the number of data in the ith zone, k,. = the mean of the permeability data in the ith zone, k . = the over-all mean of the data in the well, W = the pooled variance within zones, N = the total number of

By William L. Boyd, Eugene L. McCarthy, Laurance S. Reid

Proper control of the moisture content of natural gas is essential to reliable operation of gas transmission and distribution facilities serving northern markets. The moisture content of natural gas is usually determined by dew point measurement at the existing pressure. For any gas of constant moistrire content. the dew Point varies with the pressure. A correlation of the data of several investigators is prezented in graphical form by the authors. These data were correlated by the authors and F. M. Townsend, C. C. Tsao. M. I). Rogers. Jr.. and J. A. Porter. graduate students in chemi(.a1 engineering at the University of Oklahoma. Of articular interest are the hitherto unpublished low temperature data observed by Wickliffe Skinner. Jr., which are included in this correlation. PRESENTATION OF DATA The problem of interpreting water dew points, or saturation temperatures. of natural gas in terms of specific moisture con-tent has increased in importance during the past decade bec.arl.;e of extensive development.; in the transmission and petro-chemical phases of the natural gas industry. Virtually all gas transported to northern and eastern markets must be dehydrated to a low water vapor content to prevent hydrate formation in transmission and distribution lines and resultant interruption.; in gas deliveries. Complete dehydration is required in certain phases of tile petro-chemical industry involving low-temperature operations. It is a well-known fact that the water vapor content of pure hydrocarbon vapors and their mixtures at superatmospheric pressures cannot be predicted with accuracy by assuming validity of the ideal gas laws." Earlier interest in the general problem was concentrated on the water vapor content of pure hydrocarhons and hydrocarbon mixtures in the pressure and temperature ranges common to gas and oil producing reservoirs in order to obtain fundamental data for the improvement of production techniques and the furtherance of reservoir studies. Excellent data are published for pressures ranging from atmospheric to 10,000 psig and for temperatures ranging from 100° to 460° F4,9,10,11 and are found to be in close agreement. However, experimental data at high pressures and temperatures below 100°F are comparatively limited in scope. Experimental data in the lower temperature range have been reported by Laulhere and Briscoe.8 Deaton, et al,2,3 Hammerscllmidt,5,6 and wade," In general, the pressures employed in these investigations ranged from atmospheric to 1.000 psig while temperatures ranged from 32° to 120°F; i.e., the usual conditions encountered in gas transmission line operations, Additional data were reported by Russell, et al,12 at pressures as high as 2.000 psig and covering a rather narrow atmospheric temperature range. In 1947, Hammerschmidt published a correlation of all available data,' in which the water vapor content of gases at saturation. under high pressure and low temperature. was predicted by extrapolation. In 1948, Wickliffe Skinner. Jr.. presented data on the moisture content of a low nitrogen content gas at low temperature and at pressures ranging upward to 1.500 psia.1-3 Comparison of Skinner's experimental data with the extrapolated data of Hammerscllmidt revealed an appreciable variation in the lower temperature range. emphasizing the need for a new correlation which would rely on Skinner's data at lower temperatures. Careful scrutiny of available data suggested that presence of an appreciable quantity of nitrogen in a gas mixture may affect its saturated moisture content so that data obtained from gases with more than three per cent nitrogen were not used in this correlation. CORRELATION OF DATA Data employed in this correlation are presented in Table I. The data of Dodson and Standing.4 McKetta and Katz,9 and Olds, Sage and Lacey10 were compared and found to be in close agreement so that the data of Olds. et al, re-plotted in a more convenient form by the Humble Oil and Refining Co.,' were used for temperatures of 100°F and above. Skinner's data were used for temperatures below 40°F. Between these intermediate temperature limits, the data of Hammerschmidt.7 Wade" and extrapolated data of Olds. et al.1 were tabulated

Jan 1, 1950

By C. O. Dale, W. J. Rundle

THE upper Mississippi Valley zinc-lead area was the first major lead-producing section in the United States. The lead ore, found near the surface in crevices, was relatively pure galena that could be smelted directly into lead, at first in log hearth furnaces and later in more efficient blast type furnaces. French Canadian fur traders encouraged the Indians to mine the lead ore and showed them how to smelt it into lead that had a high value for bullets1 Nicholas Perrot found lead ore on the Mississippi River bluffs near the junction of Wisconsin and Illinois and in 1690 established a trading post on the Wisconsin side of the river opposite the present site of Dubuque, Iowa.2 Shortly after 1720 discovery of Mine La Mott in Missouri diverted considerable attention from the Upper Mississippi area. Mining continued on a desultory basis with operations concentrated in the Galena, Illinois-Dubuque, area. In 1740 at least 20 miners were at work in the Fever River area around Galena and are reported to have shipped 2500 70-lb pigs of lead to Kaskaskia in 1741." Julien Dubuque established a mining and smelting operation in 1790 near the city that bears his name and was granted sole right to exploit the mining operations on the lands of the Sauk and the Fox Indians. He is reported to have produced 30,000 70-lb pigs of lead in 1805. Following the death of Dubuque in 1810 the Indians refused to let the white miners enter their lands, and little was done on the Iowa side of the river until the Indians were removed by treaty with the United States government in 1832." Early mining was entirely for lead but as the crevices were followed down, increasing percentages of zinc sulphide and zinc carbonate were encountered and at first discarded. Later a market became available for the zinc ores, and hand jigging devices were made to separate the lead, the zinc, and the rock or waste materials. The first record of zinc production from the area is for 1860. Production of zinc passed that of lead before 1900, reached a peak of 64,000 short tons in 1917, fell off rapidly and continually to about 2000 short tons in 1938, and since 1940 has ranged from 11,000 to 19,000 short tons. Lead has been of considerably less importance since 1900, and at present only about 10 pct as much lead as zinc is produced. Practically all of the zinc ore has come from orebodies that are rather flat and wide with considerable length as compared to width. Most of the early lead came from the crevice type deposit, but present production is from the predominately flat zinc orebodies. The Graham-Snyder orebody, scene of Eagle Picher operations, is practically all zinc with little or no lead being recovered. Marcasite, present in varying amounts, makes production of finished concentrates by gravity separation impractical. Satisfactory lead and zinc concentrates have been produced since flotation was introduced in the area in 1927. An acid recovery plant was operated for about 20 years after World War I, but it has been dismantled, and no recovery of the iron sulphides in the ores of the district is being made at the present time. In June 1950 there were three companies operating mines and mills, Tri-State Zinc Co., Calumet & Hecla Consolidated Copper Co., and Eagle Picher Mining and Smelting Co. The Vinegar Hill Zinc Co. had completed a shaft at a new orebody and had started to develop the mine which will supply the Cuba City mill. The Cuba Mining Co. was holding the Andrews Mine inactive. The Dodgeville Mining Co. was not operating but was exploring for additional reserves. Several small mines were selling ore to the Eagle Picher mill. A general area map is given in Fig. 1. The Eagle Picher Mining and Smelting Co. entered the area in 1946 with an active exploration campaign. Leases on a block basis were secured for the area south from the Wisconsin-Illinois line near

Jan 1, 1953

By C. O. Dale, W. J. Rundle

THE upper Mississippi Valley zinc-lead area was the first major lead producing section in the United States. The lead ore, found near the surface in crevices, was relatively pure galena that could be smelted directly into lead, at first in log hearth furnaces and later in more efficient blast type furnaces. French Canadian fur traders encouraged the Indians to mine the lead ore and showed them how to smelt it into lead that had a high value for bullets.1 Nicholas Perrot found lead ore on the Mississippi River bluffs near the junction of Wisconsin and Illinois and in 1690 established a trading post on the Wisconsin side of the river opposite the present site of Dubuque, Iowa.2 Shortly after 1720 discovery of Mine La Mott in Missouri diverted considerable attention from the Upper Mississippi area. Mining continued on a desultory basis with operations concentrated in the Galena, Illinois-Dubuque, area. In 1740 at least 20 miners were at work in the Fever River area around Galena and are reported to have shipped 2500 70-lb pigs of lead to Kaskaskia in 1741.3 Julien Dubuque established a mining and smelting operation in 1790 near the city that bears his name 'and was granted sole right to exploit the mining operations on the lands of the Sauk and the Fox Indians. He is reported to have produced 30,000 70-lb pigs of lead in 1805. Following the death of Dubuque in 1810 the Indians refused to let the white miners enter their lands, and little was done on the Iowa side of the river until the Indians were removed by treaty with the United States government in 1832.4 Early mining was entirely for lead but as the crevices were followed down, increasing percentages of zinc sulphide and zinc carbonate were encountered and at first discarded. Later a market became available for the zinc ores, and hand jigging devices were made to separate the lead," the zinc, and the rock or waste materials. The first record of zinc production from -the area is for 1860. Production of zinc passed that of lead before 1900, reached a peak of 64,000 short tons' in 1917, fell off rapidly and continually to about 2000 short tons in 1938, and since 1940 has ranged from 11,000 to 19,000 short tons. Lead has been of considerably less importance since 1900, and at present only about 10 pct as much lead as zinc is produced. Practically all of the zinc ore has come from orebodies that are rather flat and wide with, considerable length as compared to width. Most of the early lead came from the crevice type deposit, but present production is from the predominately flat zinc orebodies. The Graham-Snyder orebody, scene of Eagle Picher operations, is practically all zinc with little or no lead being recovered. Marcasite, present in varying amounts, makes production of finished concentrates by gravity separation impractical. Satisfactory lead and zinc concentrates have been produced since flotation was introduced in the area in 1927. An acid recovery plant was operated for about 20 years after World War I, but it has been dismantled, and no recovery of the iron sulphides in the ores of the district is being made at the present time. In June 1950 there were three companies operating mines and mills, Tri-State Zinc Co., Calumet & Hecla Consolidated Copper Co., and Eagle Picher Mining and Smelting Co. The Vinegar Hill Zinc Co. had completed a shaft at a new orebody and had started to develop the mine which will supply the Cuba City mill. The Cuba Mining Co. was holding the Andrews Mine inactive. The Dodgeville Mining Co. was not operating but was exploring for additional reserves. Several small mines were selling ore to the Eagle Picher mill. A general area map is given in Fig. 1. The Eagle Picher Mining and Smelting Co. entered the area in 1946 with an active exploration campaign. Leases on a block basis were secured for the area south from the Wisconsin-Illinois line near

Jan 1, 1952

By Strathmore R. B. Cooke

On the basis of experiments conducted on quartz using a bubble pick-up method, it was shown in an earlier paper1 that this mineral will preferentially adsorb hydrogen, calcium, or sodium ions, depending on the relative concentrations of those ions in the solution in which the quartz is immersed. For quartz particles ranging in size from 0.2 to 1 mm, it was demonstrated that the concentration of calcium in solution (assumed present as ions) necessary to completely activate quartz for flotation is given by the expression: Ca++ = [H+] X 106 + [Na+] X 10-3 In this expression, ionic concentrations are in mols per liter. It was further shown that the pick-up method is apparently more sensitive to changes in reagent concentration than the standard captive-bubble method, and that induction times are apparently much reduced. Since completing these earlier tests, a new type of cell has been constructed; this is shown in Fig 1 and 2. In Fig 1, A is a ground joint, B is a central tube reaching to within a few millimeters of the bottom of the cell, and C is a stopper which may be removed for reagent addition, and serves the further purpose of excluding carbon dioxide from the air during the test. The entire cell is constructed of Pyrex glass, and does not give the trouble experienced with the earlier cell, in which activating ions were released from the glass at high pH values. The only critical factor in the construction of the cell is the clearance between the central tube and the bottom of the cell. This clearance should be sufficiently small that the bubble can be pressed directly on the mineral grains lying on the bottom of the cell. In the pick-up tests to be described in this paper, the reagents used were all of C. P. grade, except the sodium oleate, which was Merck's "neutral powder." The quartz employed was water-clear vein quartz, sized on screens, cleaned with both hydrochloric acid and sodium hydroxide, and given a thorough final washing with distilled water. Experimental procedures were the same as described in the earlier paper. EFFECT OF SIZE OF QUARTZ ON PICK-UP To ascertain the effect of particle size on the adsorption of calcium ions, the quartz was sized from minus 14 plus 20 mesh through the intervening screen sizes to minus 270 mesh plus 400 mesh. Each size was thoroughly cleaned, and then tested in the cell at different calcium chloride and sodium hydroxide concentrations, and at a constant sodium oleate concentration of 20 mg per liter. All particles, within the size range given, exhibited complete pick-up within the curve expressed by the equation above. This presumably means, when the conditions imposed by the equation are satisfied, that this maximum is independent of particle size. However, it was found that the range through which partial particle pick-up occurred progressively broadened as particle size decreased. This is shown in Fig 3, in which curves B, C, and D show the limits at which pick-up just commences (as the pH is increased) for particles of minus 14 plus 28 mesh, minus 65 plus 100 mesh, and minus 270 plus 400 mesh size, respectively. These results indicate that for satisfactory activation, at any given pH, a lower calcium ion concentration is required for fine particles than for coarse particles. EFFECT OF HIGH ALKALINITY ON PICK-UP At calcium concentrations of between 1 and 10 mg per liter, and at high alka-linities, it was noticed that pick-up ceased as soon as calcium hydroxide commenced to precipitate. This effect was investigated at other calcium concentrations, with the same results. Solutions of calcium chloride, containing 105, 104, l03, and l02 mg of calcium per liter were made alkaline with sodium hydroxide until calcium hydroxide just started to precipitate, according to the following equation: CaCl² + NaOH -+ Ca(OH)2 + 2NaCl The beginning of precipitation was taken as that point at which either a faint opalescence appeared in the solution, or a Tyndall cone became ap-

Jan 1, 1950

By J. E. Papin

MORENCI ores contain as an average about 0.015 pct molybdenite, MoS2. Incidental to the concentrating operations applied for the recovery of copper minerals, approximately two-thirds of the molybdenite is floated and appears in the final copper concentrate. The economic importance of molybdenum and the success achieved in its recovery in milling operations elsewhere encouraged research directed toward its recovery in marketable form. Several procedures are utilized to effect separation of molybdenite from copper and iron sulphides, but only two have wide application. In the better known of these two methods the molybdenite in the copper concentrate is depressed in a flotation operation using soluble starch as the depressant. The tailing from this treatment thus becomes a low-grade molybdenite concentrate which after thickening, filtering, and low temperature roasting is repulped with water and subjected to additional flotation steps for recovery of the molybdenite. A thio-phos-phate collecting agent is employed in flotation of the copper minerals. Presumably the stability and lasting effects of that collector type necessitated the practice of depressing molybdenite followed by subsequent roasting to insure elimination of copper and iron sulphides from final molybdenite concentrate. The second important method of recovery is applied at various southwestern plants. In those plants xanthates are in use as copper collecting agents. It has been found that xanthates are relatively unstable and their collecting power for sulphides is destroyed by very simple procedures. The method generally applied is to subject the xanthate-acti-vated concentrate, in the form of a pulp, to prolonged heating using steam as the heating medium. A concentrate thus treated may then be subjected to a flotation operation for recovery of the molybdenite, using mineral oil collecting agents, and the copper and iron sulphides will be depressed. Neither of the above processes was applicable to Morenci concentrates. Morenci uses a thio-phos-phate collecting agent in its flotation operation. However, it was established that soluble starch was not a depressant for Morenci molybdenite; therefore the first step in that process was not applicable. AS would be expected the copper and iron sulphides in Morenci concentrates could not be depressed by the heating methods used on xanthate activated concentrates. Low-temperature roasting did eliminate the effects of the thio-phosphate but had to be rejected as uneconomic in as much as the tonnage of concentrates involved was too great. Preliminary research having established the above facts, it became obvious that new methods would have to be developed for the recovery of molybdenite from Morenci concentrates. A broad laboratory investigation was initiated along two general lines, both directed toward depression of copper and iron sulphides. One approach was through the use of water-soluble oxidizing agents with the objective of destroying the thio-phosphate in the concentrate and thus eliminating its collector effect. The other approach involved the use of soluble sulphides, which have long been recognized as sulphide depressants. These efforts resulted in laboratory processes which showed promise. Research was then extended to a pilot plant having a daily capacity of several tons of concentrate. This operation soon demonstrated that soluble sulphide—basically sodium polysulphide—when applied to a pulp made slightly acid with sulphuric acid, proved to be the most effective sulphide depressant. The rate of concentrate treatment in this plant and the low molybdenite content of the concentrate prevented carrying the process to a conclusion, namely, production of molybdenite concentrate of marketable grade. Laboratory treatment of pilot plant concentrate indicated that a marketable product could be made from it, given a tonnage suited to available cleaner flotation machine capacity. Accordingly, the molybdenite plant was enlarged to permit treatment of the total copper concentrate from the extension plant, maximum of approximately 800 tons per day. In the enlarged plant a major difficulty was encountered. No criteria could be established as a basis for control of the process. In the early stages of flotation, because of the small amount of molybdenite involved, there was no visual evidence of the relative floatability of molybdenite with respect to

Jan 1, 1956

By J. C. Purcupile, R. L. Morris

Present methods of roof support require that first a hole must be drilled, and then a toggle belt or grouted resin bolt be set. Automating this system is very difficult. Therefore, we decided to develop a system which would insert pins into the roof in one step. We think that this type of system can be automated much easier. It would then speed up the roof bolting operation and eliminate the need for a man to work under unsupported roof, making roof bolting a much safer job. There are two ways that a pin can be inserted in a mine roof without drilling a hole. It can either be pushed in, using a very high, constant axial force; or it can be hammered in like a nail. The former method involves very high compressive stresses in the pin, and requires complicated machinery to prevent buckling. Since hammering the pin does not require high axial forces buckling is not as likely. Also, experiments have shown that there are no problems associated with getting the pin to go in straight. We chose to develop the second method. While investigating methods for hammering a pin into a mine roof, we learned that piles were being driven 10 times faster than by conventional methods with a method called "orboresonance." We decided to turn the pile driver upside down, and drive pins up into the mine roof. Sonic Pin Setting Machine The machine consists of a steel bar supported at the nodal points of its first mode of transverse vibration (Fig. 1). A mechanical oscillator, driven by a hydraulic motor, is attached to one end of the bar. This provides the forcing function for the vibrations in the bar. The hydraulic power is supplied by a portable unit. When the forcing function is near the natural frequency of the bar, large amplitudes of vibration occur and require little power input. An anvil is attached to the bar at a point of maximum deflection, providing a means for striking the pin. Since the bar is being driven near its natural frequency, it takes little energy to accelerate the bar back and forth between each hit. This means that most of the power being put into the bar is going into driving the pin. This is unlike conventional pile drivers which use most of their power to accelerate the hammer back and forth between hits. The resonant frequency of the bar we used is about 250 Hz, which is in the audible range, hence the term "sonic pin setting machine." The system benefits are: (1) elimination of the two-step drilling and pin-setting roof bolt operation, (2) increased speed, (3) more intimate contact between the formation and bolts, (4) less expensive roof bolts, (5) higher integrity, (6) potential for automatic operation. Modifications During the months from Oct. to Dec. 12, 1974, under U.S. Bureau of Mines grant No. 00144104, "Projects on Coal Extraction," six members of the senior class at Carnegie Mellon University designed and built the experimental machine. They drove three pins into the Safety Research coal mine at Bruceton, Pa., and determined that the project was worth continuing. In January 1975 we began making modifications and repairs in preparation for a more extensive program of testing at Bruceton. The major items were: 1) Design and build a collapsible pin guide to keep the pin parallel with the line of travel of the lift table. 2) Move the anvil to the center of the bar. 3) Refinish the nodal holes and install new node pins and bushings. 4) Realign the bar in its frame. 5) Install a relief valve on the lift unit to control the upward force. The upward force that we exert on the pin is relatively small, so there is no tendency for it to buckle. This means that all we need to do is hold the pin straight during its initial penetration. This allows the bushing to travel up and down in a straight line. At the beginning of penetration, the bushing

Jan 1, 1977

By R. McNeill, D. E. Weiss, E. A. Swinton

In a continuous countercurrent exchange process, an alteration in any one of the operating conditions has a complex effect on the others, which can only be predicted by employing the transfer unit or the theoretical stage theory on a basis of trial and error. A simple method is described for illustrating diagrammatically the behavior of a counter-current system, the equations being simplified by means of a concept the maximum hypothetical exchange performance. An example based on a typical metallurgical system is given, in which a divalent metal is recovered from a dilute solution, the resin being regenerated continuously by a monovalent ion. Useful conclusions are drawn from a study of the theory. Practical methods for performing continuous ion exchange are discussed, and the development of equipment based on modified ore dressing jigs is described. A swinging sieve jig contactor is evaluated experimentally. DURING the last decade, the new synthetic ion exchange resins have been applied extensively in industries outside the field of water treatment, but there is no record of a continuous counter-current process operating on an industrial scale. Attempts have been made to devise a satisfactory process but many problems remain to be solved. The basic principles of continuous processes will be outlined, as well as the major problems in their operation and the progress made in the CSIRO laboratories toward the development of satisfactory industrial techniques. In the metallurgical field ion exchange resins can be used for various applications such as the recovery and concentration of valuable metals from mine waters,' the regeneration of pickling and plating liquors," the prevention of pollution by waste effluents and the recovery of the constituents from them," and the purification of valuable metals such as the rare earths by chromatographic fractionation on columns of ion exchange resins.7,8 . Turther applications undoubtedly will be found in the field of hydrometallurgy where the use of ion exchange resins would enable direct extraction of the desired metal ion from the filtered leach liquor or the leach pulp. For example, an ion exchange process has been described recently for the extraction of gold from a cyanide leach pulp." A continuous process would have advantages in many applications over the usual process employing a fixed bed and intermittent cycle. In a recovery process, it would yield a product stream of steady purity and concentration, it would waste less water in rinsing, and if the contacting apparatus were efficient less resin would be used, since each portion of the resin would be cycled as soon as it was loaded instead of lying idle until the whole bed was ready for regeneration. A very major advantage is that it would be simpler to control automatically. It is probable that continuous operation will be the key for really large scale applications of ion exchange. The flow sheet of a continuous ion exchange recovery-concentration process is illustrated diagrammatically in Fig. 1. Dilute liquor containing the valuable ion flows through the stripping section countercurrently to a moving bed of resin and leaves after a final contact with freshly regenerated resin. The resin leaves the unit almost in equilibrium with the incoming liquor and then flows to the regenerating unit where it is treated by a slow countercurrent flow of concentrated regenerant solution. The adsorbed ion is displaced from the resin and appears in the concentrated product stream. The resin then must pass through a rinse unit or section where regenerant entrained by the resin is washed back into the regeneration section by water. The regenerated and washed resin is then recycled back to the stripping section. I. Theoretical Operating Behavior of Continuous Ion Exchange Stripping System The simple theory of continuous ion exchange is analogous to that of solvent extraction and other diffusional transfer operations and is governed by the equilibrium relationship, the mass balance, the rates of mass transfer, and the contacting efficiency of the unit. Equilibrium Relationship—The relative affinity of two ions A and B, for a particular resin immersed in their solution, can be expressed by plotting compositions of the solution against compositions which exist in resin in equilibrium with those solutions, i.e. C/Co vs q/a where C, is the total normality of the solution, C is the normality of ion A in the solution, a is the total exchange capacity of the resin in gram equivalents

Jan 1, 1956

By AIME AIME

PROBABLY no man has been of greater service to the Institute and has kept more in the background than Henry Krumb. A Vice-President continuously) for the last eleven years, apparently neither his picture nor a biographical sketch ever have adorned these pages and were he forewarned in the present instance he would order us to "forget it." He is a Columbia School of Mines man, Class of '98. He worked underground at Rossland, B. C. for a time, then for a year and a half a. chief engineer of the famous Camp Bird at Ouray, Colo. For three wars he was examining engineer for the Guggenheims and since 1901 has been an independent consulting engineer with experience throughout the Americas.

Jan 1, 1939

By Glenn S. Wyman

CHUQUICAMATA, located in northern Chile in the Province of Antofagasta, is on the western slope of the Andes at an elevation of 9500 ft. Because of its position on the eastern edge of the Atacama Desert, the climate is extremely arid with practically no precipitation, either rain or snow. All primary blasting in the open-pit mine at Chuquicamata is done by the churn drill, blasthole method. Since 1915, when the first tonnages of importance were removed from the open pit, there have been many changes in the blasting practice, but no clear-cut rules of method and procedure have been devised for application to the mine as a whole. One general fact stands out: both the ore and waste rock at Chuquicamata are difficult to break satisfactorily for the most efficient operation of power shovels. Numerous experiments have been made in an effort to improve the breakage and thereby increase the shovel efficiency. Holes of different diameter have been drilled, the length of toe and spacing of holes have been varied, and several types of explosives have been used. Early blasting was done by the tunnel method. The banks were high, generally 30 m, requiring the use of large charges of black powder, detonated by electric blasting caps. Large tonnages were broken at comparatively low cost, but the method left such a large proportion of oversize material for secondary blasting that satisfactory shovel operation was practically impossible. Railroad-type steam and electric shovels then in service proved unequal to the task of efficiently handling the large proportion of oversize material produced. The clean-up of high banks proved to be dangerous and expensive as large quantities of explosive were consumed in dressing these banks, and from time to time the shovels were damaged by rock slides. As early as 1923 the high benches were divided, and a standard height of 12 m was selected for the development of new benches. The recently acquired Bucyrus-Erie 550-B shovel, with its greater radius of operation compared to the Bucyrus-Erie 320-B formerly used for bench development, allowed the bench height to be increased to 16 m. Churn drill, blasthole shooting proved to be successful, and tunnel blasts were limited to certain locations where development existed or natural ground conditions made the method more attractive than the use of churn drill holes. Liquid oxygen explosive and black powder were used along with dynamite of various grades in blast-hole loading up to early 1937. Liquid oxygen and black powder were discontinued because they were more difficult to handle due to their sensitivity to fire or sparks in the extremely dry climate. At present ammonium nitrate dynamite is favored because of its superior handling qualities and its adaptability to the dry condition found in 90 pct of the mine. In wet holes, which are found only in the lowest bench of the pit and account for the remaining 10 pct of the ground to be broken, Nitramon in 8x24-in. cans, or ammonium nitrate dynamite packed in 8x24-in. paper cartridges, is being used. This latter explosive, which is protected by a special antiwetting agent that makes the cartridges resistant to water for about 24 hr, currently is considered the best available for the work and is preferred over Nitramon. Early churn drill hole shots detonated by electric blasting caps, one in each hole, gave trouble because of misfires caused by the improper balance of resistance in the electrical circuits. Primarily, it was of vital importance to effect an absolute balance of resistance in these circuits, the undertaking and completion of which invariably caused delays in the shooting schedule. Misfires resulting from the improper balance of electrical circuits, or from any other cause, were extremely hazardous, since holes had to be unloaded or fired by the insertion of another detonator. The advent of cordeau, later followed by primacord, corrected this particular difficulty and therefore reduced the possibility of missed holes. After much experimentation, the blasting practice evolved into single row, multihole shots, with the holes spaced 4.5 to 5 m center to center in a row 7.5 to 8 m back from the toe. Sucti shots were fired from either end by electric blasting caps attached to the main trunk lines of cordeau or primacord. The detonating speed of cordeau or primacord gave the practical effect of firing all holes instantaneously. Double row and multirow blasts, fired instantaneously with cordeau or primacord, proved to be unsatisfactory in the type of rock found at Chuquica-

Jan 1, 1953

By Glenn S. Wyman

CHUQUICAMATA, located in northern Chile in the Province of Antofagasta, is on the western slope of the Andes at an elevation of 9500 ft. Because of its position on the eastern edge of the Atacama Desert, the climate is extremely arid with practically no precipitation, either rain or snow. All primary blasting in the open-pit mine at Chuquicamata is done by the churn drill, blasthole method. Since 1915, when the first tonnages of importance were removed from the open pit, there have been many changes in the blasting practice, but no clear-cut rules of method and procedure have been devised for application to the mine as a whole. One general fact stands out: both the ore and waste rock at Chuquicamata are difficult to break satisfactorily for the most efficient operation of power shovels. Numerous experiments have been made in an effort to improve the breakage and thereby increase the shovel efficiency. Holes of different diameter have been drilled, the length of toe and spacing of holes have been varied, and several types of explosives have been used. Early blasting was done by the tunnel method. The banks were high, generally 30 m, requiring the use of large charges of black powder, detonated by electric blasting caps. Large tonnages were broken at comparatively low cost, but the method left such a large proportion of oversize material for secondary blasting that satisfactory shovel operation was practically impossible. Railroad-type steam and electric shovels then in service proved unequal to the task of efficiently handling the large proportion of oversize material produced. The clean-up of high banks proved to be dangerous and expensive as large quantities of explosive were consumed in dressing these banks, and from time to time the shovels were damaged by rock slides. As early as 1923 the high benches were divided, and a standard height of 12 m was selected for the development of new benches. The recently acquired Bucyrus-Erie 550-B shovel, with its greater radius of operation compared to the Bucyrus-Erie 320-B formerly used for bench development, allowed the bench height to be increased to 16 m. Churn drill, blasthole shooting proved to be successful, and tunnel blasts were limited to certain locations where development existed or natural ground conditions made the method more attractive than the use of churn drill holes. Liquid oxygen explosive and black powder were used along with dynamite of various grades in blast-hole loading up to early 1937. Liquid oxygen and black powder were discontinued because they were more difficult to handle due to their sensitivity to fire or sparks in the extremely dry climate. At present ammonium nitrate dynamite is favored because of its superior handling qualities and its adaptability to the dry condition found in 90 pct of the mine. In wet holes, which are found only in the lowest bench of the pit and account for the remaining 10 pct of the ground to be broken, Nitramon in 8x24-in. cans, or ammonium nitrate dynamite packed in 8x24-in. paper cartridges, is being used. This latter explosive, which is protected by a special antiwetting agent that makes the cartridges resistant to water for about 24 hr, currently is considered the best available for the work and is preferred over Nitramon. Early churn drill hole shots detonated by electric blasting caps, one in each hole, gave trouble because of misfires caused by the improper balance of resistance in the electrical circuits. Primarily, it was of vital importance to effect an absolute balance of resistance in these circuits, the undertaking and completion of which invariably caused delays in the shooting schedule. Misfires resulting from the improper balance of electrical circuits, or from any other cause, were extremely hazardous, since holes had to be unloaded or fired by the insertion of another detonator. The advent of cordeau, later followed by primacord, corrected this particular difficulty and therefore reduced the possibility of missed holes. After much experimentation, the blasting practice evolved into single row, multihole shots, with the holes spaced 4.5 to 5 m center to center in a row 7.5 to 8 m back from the toe. Sucti shots were fired from either end by electric blasting caps attached to the main trunk lines of cordeau or primacord. The detonating speed of cordeau or primacord gave the practical effect of firing all holes instantaneously. Double row and multirow blasts, fired instantaneously with cordeau or primacord, proved to be unsatisfactory in the type of rock found at Chuquica-

Jan 1, 1953

By H. R. Froning, R. O. Leach

In wertability alteration flooding, a chemical agent is rnoved through a reservoir by the flood water to increase oil recovery by decreasing the degree of wetting of the rock by the oil. Substantial amounts of the chemical may be lost during movement through the reservoir. The extent of the loss, and therefore the economics of the process, depends in some cases on factors which are difficult to reproduce in the laboratory. Therefore, a short-duration, low-cost field test method is needed to permit evaluation of chemical requirements under actual field conditions. This paper describes a small scale rest conducted at a single well for measuring chemical requirements, thereby giving a more reliable evaluation of this important factor in the applicability and economics of the process. In the rest a small wafer slug containing the chemical agent and a nonadsorbed tracer is displaced into the reservoir by a known volume of wafer. The well is then placed on producrion. Chemical loss per barrel of pore volume contacted is calculated from [he fractional recoveries of the agent tested and the nonadsorbed tracer. The method has been used to determine within the actual reservoirs the chemical requirements for both a sandstone and a dolomite reservoir. Several chemical agents are potentially available for wettability alteration flooding, although none is universally applicable. For some applications of the method, chemical costs per barrel of additional oil recovered can be substantially less than one dollar. INTRODUCTION Wettability alteration flooding provides a means of increasing oil recovery from reservoirs by decreasing the degree of wetting of the rock by the oil and increasing the displacement efficiency of the flood water. Earlier studies demonstrated a relationship between oil recovery during waterflooding and the degree of wetting of a rock surface by an oil. The application of wettability alteration flooding to the Harrisburg field of Nebraska provided a field test' of this recovery process. Subsequently, additional laboratory and field tests have developed additional procedures for evaluating wettability alteration flooding, and have indicated where the process may be applicable. Applicability of this process to specific reservoirs is determined by a progression of tests to determine sus- ceptibility of the reservoir to alteration of its wettability, to indicate the degree of recovery improvement and to estimate the amount of chemical required to process the reservoir. The economics of applying improved oil recovery processes depends not only upon the degree of improvement in oil recovery achievable by the process but also upon the process costs and the timing of the income and the investment. Emphasis in this paper is on the expenditure aspects of the process. The work reported in this paper indicates that the chemical investments required for wettability alteration flooding are substantial. For evaluating the economics of a potential flooding application it is imperative that a sound estimate of the chemical requirements be made for the reservoir. Generally, true reservoir conditions are not adequately simulated in laboratory chemical propagation tests. Because of wide well spacings, many years might be required to obtain chemical propagation data from conventional pilots or inter-well tests. Consequently, n short-duration, low-cost method is needed to determine chemical requirements in the field. The potential applicability of wettability alteration flooding is discussed, as well as the economics of wettability alteration with respect to the inherent and imposed restrictions on the timing of income and investments. DETERMINATION OF CHEMICAL REQUIREMENTS In the process of moving a chemical bank through reservoir rock, some of the chemical agent lags too far behind the flood front to be effective or is otherwise lost to the reservoir system. The extent to which these losses occur overshadows the reductions in chemical concentration due to diffusion and to mixing with the reservoir fluids. Experience indicates that almost without exception, chemicals which induce a wetting change undergo either sorption reactions or chemical reactions with mineral constituents of the pore surfaces. Other reactions may occur between the added chemical and the reservoir oil and water. Even in limiting consideration to reactions of the relatively inexpensive inorganic salts, bases and acids, the reactions may be exceedingly complex. Reservoir pore surfaces consist of more than silica in sandstone reservoirs, and more than calcite or dolomite in carbonate reservoirs. Many mineral species are present, each exhibiting specific tendencies to react with an injected chemical. The reactions which occur can consume enough of the agent to have an important effect on economics. These reactions can cause a change in pH of the chemical bank, or may remove some of the active chemical by precipitation, adsorption or reaction to form a new chemical which may or may not be effective in

By Glenn S. Wyman

CHUQUICAMATA, located in northern Chile in the Province of Antofagasta, is on the western slope of the Andes at an elevation of 9500 ft. Because of its position on the eastern edge of the Atacama Desert, the climate is extremely arid with practically no precipitation, either rain or snow. All primary blasting in the open-pit mine at Chuquicamata is done by the churn drill, blasthole method. Since 1915; when the first tonnages of importance were removed from the open pit, there have been many changes in the blasting practice, but no clear-cut rules of method and procedure have been devised for application to the mine as a whole. One general fact stands out: both the ore and waste rock at Chuquicamata are difficult to break satisfactorily for the most efficient operation of power shovels. Numerous experiments have been made in an effort to improve the breakage and thereby increase the shovel efficiency. Holes of different diameter have been drilled, the length of toe and spacing of holes have been varied, and several types of explosives have been used. Early blasting was done by the tunnel method. The banks were high, generally 30 m, requiring the use of large charges of black powder, detonated by electric blasting caps: Large tonnages were broken at comparatively low cost, but the method left such a large proportion of oversize material for secondary blasting that satisfactory shovel operation was practically impossible: Railroad-type steam and electric shovels then in service proved unequal to the task of efficiently handling the large proportion of oversize material produced. The clean-up of high banks proved to be dangerous and expensive as large quantities of explosive were consumed in dressing these banks, and from time to time the shovels were damaged by rock slides. As early as 1923 the high benches were divided, and a standard height of 12 m was selected for the development of new benches. The recently acquired Bucyrus-Erie 550-B shovel, with its greater radius of operation compared to the Bucyrus-Erie 320-B formerly used for bench development, allowed the bench height to be increased to 16 m. Churn drill, blasthole shooting proved to be successful, and tunnel blasts were limited to certain locations where development existed or natural ground conditions made the method more attractive than the use of churn-drill holes. Liquid oxygen explosive and black powder were used along with dynamite of various grades in blasthole loading up to early 1937. Liquid oxygen and black powder were discontinued because they were more difficult to handle due to their sensitivity to fire or sparks in the extremely dry climate. At present ammonium nitrate dynamite is favored because of its superior handling qualities and its adaptability to the dry condition found in 90 pct of the mine. In wet holes, which are found only in the lowest bench of the pit and account for the remaining 10 pct of the ground to be broken, Nitramon in 8x24-in. cans, or ammonium nitrate dynamite packed in 8x24-in. paper cartridges, is being used. This latter explosive, which is protected by a special antiwetting agent that makes the cartridges resistant to water for about 24 hr, currently is considered the best available for the work and is preferred over Nitramon. Early churn drill hole shots detonated' by electric blasting caps, one in each hole, gave trouble because of misfires caused by the improper balance of resistance in the electrical circuits. Primarily, it was of vital importance to effect an absolute balance of resistance in these circuits, the undertaking and completion of which invariably caused delays in the shooting schedule. Misfires resulting from the improper balance of electrical circuits, or from any other cause, were extremely hazardous, since holes had to be unloaded or fired by the insertion of another detonator. The advent of cordeau, later followed by primacord, corrected this particular difficulty and therefore reduced the possibility of missed holes. After much experimentation, the blasting practice evolved into single row, multihole shots, with the holes spaced 4.5 to 5 m center to center in a row 7.5 to 8 m back from the toe. Such shots were fired from either end .by electric blasting caps attached to the main trunk lines of cordeau or primacord. The detonating speed of cordeau or primacord gave the practical effect of firing all holes instantaneously. Double row and multirow blasts, fired instantaneously with cordeau or primacord, proved to be unsatisfactory in the type of rock found at Chuquica-

Jan 1, 1952

By R. D. Lord

The search for the best practical means of storing sulfide bearing tailings, where there is no residual excess of carbonate material is discussed in this paper• Usually the sulfide content decomposes, with the aid of bacterial action, and the resulting sulfuric acid escapes, along with any heavy-metal solutes, through embankments that are usually porous to some degree• The problem is typified in the tailings of the uranium operations of Elliot Lake, Ont., where mining started some 20 years ago• The approach to tailings disposal paralleled the practice for other hydrometallurgical plants treating gold and base-metal ores• Impoundment areas were designed to retain solids, and a clear and neutral overflow was considered satisfactory practice• Now experience has shown that these areas, some of which have been idle for over a dozen years, release acids in seepage and overflows to an unacceptable degree• To protect natural water courses, neutralizing plants are operated wherever required• Lime slurry is fed continuously into the tailings outflows in a quantity sufficient to raise the pH to 8•5 and precipitate heavy metals that may be in solution• The objection to this procedure is that the plants will require servicing indefinitely, unless a better remedy is found• The problem differs only slightly from that common to base-metal concentrators in that here the ore has been leached with sulfuric acid for the recovery of uranium• Any native content of calcareous material has been digested, and only that added for final neutralization is available to maintain a pH unfavorable to bacterial activity• Chemical oxidation slowly lowers the pH and when this reaches a level of 4•5 or less, bacteria become active and greatly accelerate the formation of acid. The bacterial process is probably at least ten times as fast as the chemical oxidation• Location and Processing The operations referred to, uranium and one copper mine, are located at approximately 46°N and 82°W longitude• This is typical Canadian Shield country, a land of lakes, deeply glaciated and rocky, with sparse soil which supports mixed forest cover• Drainage is to Lake Huron, 25 miles to the south• Average temperature is 45°F, ranging from -40° to +95°F• Annual precipitation is 38 in•, about half of which is snow• The ore is Precambrian, quartz-pebble conglomerate, with mineralization in the matrix• From 5 to 10% pyrite is present• All known means of pre-concentration have been tested, but a bulk sulfuric acid leach has proved the most efficient. Tailings have from the outset been neutralized before release• Current practice is to add ground limestone to bring the pH to 4•5, and then lime to raise the value to 10•5• Environmental regulations have recently been increased and the foregoing meets the new standards• Separate measures are taken to precipitate radium• Remedial Measures Since the outstanding environmental problem is the oxidation of pyrite by bacterial action, the solution is to contain the products, or arrest the process• Given the ambient temperature, favorable half of the time, four items are essential to the activity• 1) Pyrite• 2) Moisture pH < 4•5. 3) Oxygen• 4) Bacteria• Removing any one of these out of the range of tolerance will bring the reactions under control• A variety of proposals considered, and a number tested for the arrest of the process, are: (a) render embankments impermeable, (b) provide an impermeable cover, (c) cover with an oxygen absorbing layer, (d) provide a vegetative cover, (e) flood the site, (f) remove pyrite from current tailings, (g) add excess limestone to current tailings, (h) poison the bacteria• Bank Seal-On existing impoundment areas, where the embankments are several thousand yards in length, it is believed that any program of injecting sealants can have small chance of success• However, a moisture barrier is an indicated specification for future construction, and this can be highly expensive• Surface Seal-Depending on the configuration of the deposit, the downward travel of water should be prevented, and oxygen excluded• Burying a plastic membrane just below the surface has been considered, as has the application of a liquid sealant that would penetrate the surface. The objection to these remedies is the excessive cost of dealing with large areas and the expectation of only temporary benefit as a result• Frost penetration is over 4 ft, and frost action breaks up asphalt paving and all but heavy concrete in a few years• Organic Layer-An oxygen-absorbing layer, such as bark fines from paper mills has been proposed as a surface treatment• Cultivated into the tailings such material might be expected to arrest subsurface oxidation for some years• Estimates are 100 tons per acre of bark fines, or 35 tons per acre of sawdust, and these enormous quantities do not so far give assurance of providing a long-term remedy• Vegatative Cover-Several obvious benefits would result from a good growth of grass or other vegetation on abandoned tailings• While restoring the natural green of the tract the growth would prevent wind-blown dust and reduce erosion• Subsurface oxidation should be reduced, as well as the upward movement of ground moisture as occurs in dry weather. To this end, considerable research and field testing has been carried out to arrive at a formula - a prescription which will provide a self-sustaining growth on the tailings surface, or at least one that would survive with reasonable maintenance attention. Many test plots have been run with different combinations of surface treatment and seed mixtures. Generally, by addition and close cultivation of limestone, lime, and fertilizers, technical success has been demonstrated• Plants with a high tolerance for acid soil seem the more hardy, and a pH above 3 is indicated so that nutrients can be absorbed• Recommendations are for 12 to 15 tons of

Jan 1, 1977

By Carlos G. Valenzuela

The value generally accepted for stacking-fault energy, of copper has been approximately 40 ergs per sq cm based on Fullman&apos;s2 value for the coherent twin-boundary energy and the assumption that is twice the twin boundary energy However, Thornton and coworkers8 determined that the lower limit of for copper is approximately 60 ergs per sq cm, based on measurements of dislocation-node radii. They suggested that the assumption of = 2t is invalid. It is the purpose of this note to present results which indicate that the values of obtained by measurement of twin-grain boundary intersections and the values obtained by measurement of dislocation-node radii are actually compatible. The most reliable values for of the a brasses, based on electron-microscope observation of dislocation nodes, are probably those values reported by Smallman and Green6 who applied the Siems correction (1961) to data obtained by Howie and Swann.3 These values shown in Fig. 1 give an extrapolated value of approximately 70 ergs per sq cm for pure copper. In the present work, Fullman&apos;s method for determining the stacking-fault energy of copper was extended to the a brasses. High-purity copper (99.999 pet) and high-purity zinc (99.999 pet) were mixed in proportioned amounts and melted in sealed and evacuated quartz tubes. The alloys were homogenized for a week at 750°C and chemically analyzed by means of X-ray fluorescence. These alloys, along with a specimen of pure copper, were rolled to 98 pet reduction and a thickness of 0.009 in., resealed, and annealed for 40 hr at 715°C. Values for the ratio of twin-boundary energy to the grain-boundary energy, it gb, were obtained from measurements of dihedral angles formed at the intersections of twin boundaries and grain boundaries by using Fullman&apos;s2 mechanical analogy of surface tensions acting at the intersection of an annealing twin and a grain boundary. The angles between the twin traces and the grain boundaries were measured for a large number of twin-grain boundary intersections at a magnification of X500 using a rotating mechanical stage on a Reichert metallograph. The rotary stage has a calibrated angle scale with vernier so that measurements can be made to 0.1 deg. The mean twin-grain boundary energy and standard deviation were calculated for copper and each of the brasses by the use of an IBM 7072 computer. It was found that the mean value of gb stabilized after approximately 100 twin-angle measurements. This was determined by plotting the mean against the number of angles measured. Since the measured angles deviate from true dihedral angles, a correction factor for grain orientation, discussed below, was applied. The stacking-fault energy was then calculated from these values and the values of grain-boundary energy derived by Taylor.7 The data obtained are tabulated in Table I. Values for stacking-fault energy are plotted in Fig. 1. The value of gb for the specimen of 1.03 pet Zn is close to the value obtained by Fullman2 for OFHC (99.98 pet) copper, 0.045. It is believed that the purity of the copper affects gb significantly. The 99.999 pet purity Cu used in this investigation yielded a value of yt/ygb, 0.76, which is much higher than that obtained by Fullman. Additional evidence suggesting that the purity of

Jan 1, 1965

By A. W. Schlechten, Ernest J. Breton Jr.

Experiments on the separation of copper and zinc ions by selective action of ion exchange resins showed the carboxylic type to be more effective than the sulphonic resins. The latter demonstrated a greater capacity over a wider pH range. Data show the effectiveness of resins as a means of concentration. IN recent years the restrictions of stream pollution laws and the high price of metals have created an interest in ion exchange as a means for metal recovery. Some applications have proved successful. In Germany during World War 11, 17 tons of copper per day were recovered from rayon mill wastes by means of ion exchange resins;&apos; and for some time in this country a large ion exchange unit has been in operation for the recovery of copper from rayon waste water. The possibilities of applying ion exchange to the recovery of metals occurring in plating rinse water is particularly promising. In most of these applications only the metal being recovered occurs in the waste. The ion exchange resins act merely as a means of concentrating the metals to a point where they can be recirculated. It would be highly desirable to use ion exchange as a means of not only concentrating but also of separating metals. With the exception of the impressive separations accomplished in connection with the atomic energy program, very little has been done on metal separations.&apos; Therefore, an investigation was undertaken at the Missouri School of Mines and Metallurgy to determine if either of the two main types of ion exchange resins could be used to separate metal ions in solution. The selective removal of copper ions from a mixture of copper and zinc on carboxylic and sulphonic-type resins was investigated as a function of flow rate, pH, copper-zinc ratio, and concentration. It was shown that zinc can be separated from copper and that very large ratios of concentration can be obtained using ion exchange resins. Since ion exchange is relatively new to the field of metallurgy, a brief review of the subject will be included. Theory of Ion Exchange A comprehensive theory for ion exchange has not been developed as yet, but the mechanisms are analogous to metathetical reactions: R Na + Cu++ *=? K(SO3)2 Cu + 2Na+ R is the designation for the ion exchange resin. If a copper solution is passed over a resin bed in the sodium form, two ions of sodium will be released for every ion of copper removed. For the most part this reaction follows the laws of mass action and of electrical neutrality. Consequently, if an excess of sodium ions is passed over a bed containing copper, the reactions will be reversed, and the resin will be regenerated to its original form. A few empirical rules governing the exchange reaction have been set forth: 1—In general ions with a high valence will replace ions with a lower valence. 2—Ions having higher activity coefficients have a higher replacement potential. 3—In a series of mono-valent ions, those with the smallest radii of hydra-tion will tend to replace those having larger radii of hydration. 4—Where ions are similar in most respects, those with the higher atomic weight sometimes will take precedence. This last rule is not as definite as some of the others. These rules apply to rather dilute solutions at moderate temperatures and assume all ions to be present in about equal concentrations. Higher concentrations and temperatures may in some cases reverse the normal exchange reactions. Ion exchange materials are unique in that their efficiency increases as the concentration of the solution decreases. For many exchangers, most efficient operation is obtained at concentrations in the order of one thousandths of a percent. Most applications, though, are made in solutions containing considerably higher concentrations than this. Coste9 as shown that ion exchange resins will remove aluminum and iron effectively&apos; from solutions of up to 10 pct chromic acid. Ion Exchange Resins Ion exchange resins are insoluble, porous, resinous structures to which active groups have been attached. Active groups such as (—SO,,)- and (COO)- pick up cations; hence structures saturated with groups such as these are called cation exchangers. Structures saturated with groups such as (—NH,)&apos; which pick up anions, are referred to as anion exchangers. The resinous structure of necessity is resistant to strong acids, bases, oxidizing, and reducing agents, and most of the common organic solvents. An idea of the stability can be gaged from the fact that resins last for many years under constant use without detectable chemical or physical breakdown. The ion exchange reaction is not confined to the surface of these synthetic resins. Its porous structure permits active groups in the center of a particle as well as those on the surface to remove ions. A high capacity resin such as Amberlite IR-120 will remove up to 3.3 lb Cu per cu ft of resin. In this investigation several approaches to the problem of separating copper from zinc by ion exchange were considered. First, if a reagent could be found which would complex one of these metals and not the other, then by passing this reagent through a bed of exchanger containing copper and zinc, the

Jan 1, 1952