C'- rO •P?^//^'^ Description of Manganese Nodule Processing Activities for Environmental Studies Volume III. Processing Systems Technical Analyses August 1977 A contract repoiipbrepared for: U.S. depabtiJInt of commerce Nationa* Jceai^c and Atmospheric Administration \ Office of IfAaripe Minerals \ a o o Description of IVIanganese Nodule Processing Activities for Environmental Studies Volume III. Processing Systems Technical Analyses August 1977 Prepared by Dames and Moore Salt Lake City, Utah and EIC Corporation Newton, Massachusetts for U.S. DEPARTMENT OF COMMERCE National Oceanic and Atmospheric Administration Office of Marine Minerals Rockville, Maryland / Contract 6-35331 \ National Oceanic and Atmospheric Administration Office of Marine Minerals ABSTRACT: This contract report was prepared for NOAA's Office of Marine Minerals to bring together technological information pertinent to the proc- essing of manganese nodules mined from the sea floor and to help access the effects of manganese nodule processing activities on land and marine environments. It treats potential methods, operations, and requirements for processing and handling manganese nodules on land and at sea, including recovery of value metals, transport of manganese-nodule ore and process wastes, treatment and disposal of wastes, material and energy balances, and resource requirements. The report is available through the U.S. Department of Commerce's National Technical Infonnation Seirvice, 5285 Port Royal Rd. , Springfield, VA 22151 (telephone: 703-557-4600), in three volumes: Volume I. Processing Systems Summary Volume II. Transportation and Waste Disposal Systems Volume III. Processing Systems Technical Analyses NOTICE: The findings compiled in these reports, and interpretations expressed therein, do not necessarily represent the viewpoints of the National Oceanic and Atmospheric Administration or the United States Department of Commerce. The United States — while making this information available because of its obvious value and in the public interest — assumes no responsibility for any of the views expressed therein. The National Oceanic and Atmospheric Administration does not approve, recommend, or endorse any proprietary product or proprietary material mentioned in this publication. No reference shall be made to the National Oceanic and Atmospheric Administration that would imply — directly or indirectly — that the National Oceanic and Atmospheric Administration approves or disproves of the use of any proprietary product or proprietary material mentioned herein. XX ^ DESCRIPTION OF MANGANESE NODULE PROCESSING ACTIVITIES FOR ENVIRONMENTAL STUDIES Table of Contents - Volume III Processing Systems Technical Analyses Section Page 1,0 INTRODUCTION 1-1 1.1 General 1-1 1.2 NOAA Environmental Program 1-1 1.3 Scope of This Study 1-3 2.0 ALTERNATIVES FOR ON-SHORE PROCESSING OF MANGANESE NODULES 2-1 2.1 Basis for this Study 2-1 2.2 Processing Alternatives 2-1 2.2.1 Classification Schemes 2-1 2.2.2 The Influence of External Variables 2-2 2.2.3 Acid Sulfate Lixiviant Processes 2-4 2.2.4 Chloride (Halide) Lixiviant Processes 2-5 2.2.5 Ammoniacal Lixiviant Processes 2-6 2.2.6 Smelting 2-7 2.2.7 Miscellaneous Processes 2-8 2.2.8 Metals Separation and Recovery 2-8 2.2.8.1 Ammoniacal Solutions 2-9 2.2.8.2 Acidic Sulfate Solutions 2-9 2.2.8.3 Chloride Solutions 2-11 2.3 Selection of Likely Processes 2-12 3.0 ECONOMICS OF NODULE MINING AND PROCESSING 3-1 3.1 Introduction 3-1 3.2 Ore Reserves and Grades 3-1 3.3 Mining and Processing Technology 3-2 3.4 Capital and Operating Costs 3-3 3.5 Mining 3-5 3.6 Metal Markets and Prices 3-8 3.6.1 Nickel 3-8 3.6.2 Copper 3-10 3.6.3 Cobalt 3-10 3.6.4 Manganese 3-10 3.7 Potential Revenues and Profits 3-11 3.8 Conclusions 3-13 3.9 Economic Citations 3-13 lii Table of Contents (cont'd.) Section Page 4,0 GENERAL STUDY APPROACH AND ASSUMPTIONS 4-1 4.1 Approach 4-1 4.2 Assumptions 4-1 4.2.1 Manganese Nodules Composition 4-1 4.2.2 Selection of Potentially Toxic or Hazardous Constituents 4-4 4.2.3 Hazardous Constituents of Process Supplies 4-10 4.2.4 Processing Rate 4-12 4.2.5 Product Forms 4-12 4.2.6 By-Products 4-12 4.2.7 Energy Sources 4-12 4.2.8 Water 4-13 4.2.9 Materials and Supplies 4-13 4.2.10 Process Waste and Emission Treatment Procedures .... 4-14 4.2.11 Plant Waste Products 4-15 4.2.12 Potential Accidents with Pollutant Release 4-15 5.0 REDUCTION/ AMMONIACAL LEACH PROCESS 5-1 5.1 Process Description 5-1 5.1.1 Summary Process Description 5-1 5.1.2 Detailed Process Description 5-2 5.2 Process Alternatives 5-19 5.3 Plant Operating Summary Tables 5-20 5.4 Distribution of Selected Elements in Manganese Nodules Processing: Reduction/Ammoniacal Leach Process 5-26 5.4.1 Explanation and Qualifications 5-26 5.4.2 Distributions at Process Junctures 5-26 5.4.3 Diagrams of Attenuation Factors 5-27 5.4.4 Ultimate Fate of Toxic Elements 5-43 5.4.5 Distribution of Toxic Elements Among Waste Streams . . 5-44 5.5 Plant Support Requirements 5-46 5.6 Accidental Releases 5-46 5.6.1 Elements Pertinent to the Hazards Analysis 5-46 5.6.2 Description of Potential Accidents 5-48 5.6.3 Frequency and Magnitude of Potential Accidents 5-49 5.6.4 Gaseous Releases 5-51 5.6.5 Summary of Accidental Release Estimations 5-51 5.7 Criteria Sheets 5-52 iv i Table of Contents (cont'd.) Section Page 6.0 DETAILED DESCRIPTION - CUPRION/AMMONIACAL LEACH PROCESS 6-1 6.1 Process Description 6-1 6.1.1 Summary Process Description 6-1 6.1.2 Detailed Process Description 6-3 6.2 Process Alternatives 6-20 6.3 Plant Operating Summary Tables 6-21 6.4 Potentially Toxic Element Flows and Distribution 6-27 6.5 Plant Support Requirements 6-27 6.6 Accidental Releases 6-36 6.7 Criteria Sheets 6-37 7.0 DETAILED DESCRIPTION - HIGH TEMPERATURE SULFURIC ACID LEACH PROCESS 7-1 7.1 Process Description 7-1 7.1.1 Summary Process Description 7-1 7.1.2 Detailed Process Description 7-3 7.2 Process Alternatives 7-18 7.3 Plant Operating Summary Tables 7-23 7.4 Potentially Toxic Element Flows and Distribution 7-23 7.5 Plant Support Requirements 7-35 7.6 Accidental Releases 7-36 7.6.1 Elements Pertinent to Hazards Analysis 7-36 7.6.2 Description of Potential Accidents 7-37 7.6.3 Magnitude and Frequency of Accidental Releases 7-37 7.7 Criteria Sheets 7-38 8.0 DETAILED DESCRIPTION - REDUCTION/HYDROCHLORIC ACID LEACH PROCESS. . 8-1 8.1 Process Description 8-1 8.1.1 Summary Process Description 8-1 8.1.2 Detailed Process Description 8-4 8.2 Process Alternatives 8-21 8.3' Plant Operating Summary Tables 8-24 8.4 Potentially Toxic Element Flows and Distributions 8-26 8.5 Plant Support Requirements 8-38 8.6 Accidental Releases 8-39 8.6.1 Elements Pertinent to the Hazards Analysis 8-39 8.6.2 Description of Potential Accidents 8-41 8.6.3 Magnitude and Frequency of Accidental Releases 8-41 ■8.6.4 Summary of Reduction/Hydrochloric Acid Leach Process Hazards 8-41 8.7 Criteria Sheets 8-41 \ Table of Contents (cont'd.) Section Page 9.0 DETAILED DESCRIPTION - SMELTING PROCESS 9-1 9.1 Process Description 9-1 9.1.1 Summary Process Description 9-1 9.1.2 Detailed Process Description 9-4 9.2 Process Alternatives 9-28 9.3 Plant Operating Summary Tables 9-30 9.4 Potentially Toxic Element Flows and Distribution 9-30 9.5 Plant Support Requirements, Smelting Process 9-45 9.6 Accidental Releases 9-45 9.6.1 Elements Pertinent to Hazard Analysis 9-45 9.6.2 Description of Potential Accidents 9-46 9.6.3 Magnitude and Frequency of Accidental Releases 9-47 9.7 Criteria Sheets 9-47 10.0 AT-SEA PROCESSING 10-1 10.1 Rationale for At-Sea Processing 10-1 10.2 Process Possibilities for At-Sea Treatment 10-1 10.2.1 Beneficiation 10-1 10.2.2 Partial Processing 10-1 10.2.3 Metal Production 10-2 10.3 Technological Constraints 10-2 10.3.1 At-Sea Constraints 10-2 10.3.2 Physical Benef iciation Constraints 10-3 10.3.3 Processing Constraints 10-3 10.4 Selection and Description of Possible At-Sea Treatment Schemes 10-6 10.4.1 Partial Processing at Sea 10-6 10.4.2 Reduction/Ammonia Leach Partial Process Options .... 10-7 10.4.3 High Temperature Sulfuric Acid Leach Partial Process Options 10-14 10.4.4 Reduction/Hydrochloric Acid Leach Partial Process Options 10-16 10.4.5 Cuprion-Ammoniacal Leach Partial Process Options. . . . 10-17 10.5 Processing Vessels 10-20 10.5.1 General Physical Requirements 10-20 10.5.1.1 Processing Plant Equipment 10-20 10.5.1.2 Nodule Storage 10-20 10.5.1.3 Reagents and Supplies Storage 10-20 10.5.1.4 Steam 10-21 10.5.1.5 Electrical Power 10-21 10.5.1.6 Water Requirements 10-21 VI \ ^, ¥ Table of Contents (cont'd.) Section Page 10.5.1.7 Hotel Facilities 10-22 10.5.1.8 Fuel 10-22 10.5.1.9 Transfer to Other Vessels 10-22 10.5.1.10 Feasibility of Processing Plants on Barges . 10-22 10.5.2 Shipboard Process Plants 10-23 10.5.2.1 Reduction-Ammonia Leach 10-23 10.5.2.2 High Temperature Sulfuric Acid Process . . . 10-27 10.5.2.3 Reduction Hydrochloric Acid Leach 10-28 10.5.2.4 Cuprion Ammonia Leach Process 10-29 10.5.3 Summary and Possible Alternatives to Selected Process Vessels 10-29 10.5.3.1 Summary 10-29 10.5.3.2 Process Alternatives 10-32 10.5.3.3 Vessels Alternatives 10-33 10.6 Characterization of Land-Based Processing Plants for At-Sea Preprocessing 10-33 10.6.1 Reduction/Ammoniacal Leach Process 10-33 10.6.2 High Temperature Sulfuric Acid Leach Process 10-36 10.6.3 Reduction/Hydrochloric Acid Leach Process 10-39 10.6.4 Cuprion/ Ammonia Leach Process 10-43 11.0 WASTE TREATMENT 11-1 11.1 Waste Treatment in the Base Case 11-1 11.2 Types of Wastes 11-1 11.3 Waste Treatment Alternatives 11-5 11.4 Process Description 11-7 11.5 Impact of Waste Treatment on Process Material and Energy Balances 11-14 12.0 NOISE ASSESSMENT 12-1 12.1 Introduction 12-1 12.2 Noise Sources 12-1 12.3 Noise Levels 12-7 12.4 References Cited 12-7 13.0 GLOSSARY 13-1 14.0 BIBLIOGRAPHY 14-1 / / vii List of Tables Number Page Table 2.1 Some Nodules Compositions and their Values as Ore at Current Prices 2-3 Table 3.1 Nodule Operation Capital Requirements: Esti- mates from the Literature 3-3 Table 3.2 Nodule Operating Costs: Estimates from the Literature 3-5 Table 3.3 Nodule Mining Cost Estimates 3-6 Table 3.4 Three-Metal Nodule Processing Cost Estimates 3-7 Table 3.5 Four-Metal Nodule Mining Cost Estimates 3-9 Table 3.6 Four-Metal Nodule Processing Cost Estimates 3-9 Table 3.7 Nodule Operation Revenues 3-11 Table 3.8 Potential Profits of a Three-Metal Nodule Processing Operation 3-11 Table 3.9 Potential Profits of a Four -Metal Nodule Pro- cessing Operation 3-12 Table 4.1 Survey of Nodules Compositions 4-2 Table 4.2 Major and Value Metals in Pacific Ocean Nodules '^-5 Table 4.3 Innocuous Nonminor Elements in Pacific Ocean Nodules 4-6 Table 4.4 Minor Toxic and Hazardous Elements in Pacific Ocean Nodules 4-7 Table ^-5 Composition of Sea Water 4-8 Table 4.6 Major Categories of Elements in Manganese Nodules for Purposes of this Study 4-9 Table 4.7 Drinking Water Standards for Inorganic Substances 4-10 Table 4.8 Mean Analytical Values for Chemical Elements in Coal 4-11 Table 4.9 Trace Elements in a Missouri Limestone 4-11 Table 4.10 Assumed Coal Composition 4-13 Table 4.11 Composition of Process Materials and Supplies 4-14 Table 5.1 Plant Inputs and Outputs - Reduction/Ammoniacal Leach Process 5-21 Table 5.2 Reduction/Ammoniacal Leach Process - Services Profile 5-22 viix List of Tables (cont'd) Number Table 5.3 Table 5.4 Table 5.5 Table 5.6 Table 5.7 Table 5.8 Table 5.9 Table 6.1 Table 6.2 Table 6.3 Table 6.4 Table 6.5 Table 6.6 Table 6.7 Table 6.8 Table 7.1 Table 7.2 Page Constituents of Solid Effluent Streams - Reduction/Ammoniacal Leach Process 5-23 Composition of Liquid Effluent Streams - Reduction/Ammoniacal Leach Process 5-24 Compositions of Gaseous Effluent Streams - Reduction/Ammoniacal Leach Process 5-25 Reduction of Potentially Toxic Elements in Manganese Nodule Processing - Reduction/ Ammonical Leach Process 5-28 Summary Flows and Distribution of Potentially Toxic Elements - Reduction/Ammoniacal Leach Process 5-43 Summary of Waste Streams - Reduction/Ammoniacal Leach Process 5-44 Concentrations of Toxic Elements in Waste Streams - Reduction/Ammoniacal Leach Process 5-45 Plant Inputs and Outputs - Cuprion/Ammoniacal Leach Process 6-22 Cuprion/Ammoniacal Leach Process - Services Profile 6-23 Constituents of Solid Effluent Streams - Cuprion/ Ammoniacal Leach Process 6-24 Compositions of Liquid Effluent Streams - Cuprion/ Ammoniacal Leach Process 6-25 Composition of Gaseous Effluent Streams - Cuprion/ Ammoniacal Leach Process 6-26 Summary Flows and Distribution of Potentially Toxic Elements - Cuprion/Ammoniacal Leach Process 6-33 Summary of Waste Streams - Cuprion/Ammoniacal Leach Process 6-34 Concentration of Toxic Elements in Waste Streams - Cuprion/Ammoniacal Leach Process 6-35 Plant Inputs and Outputs - High Temperature Sul- furic Acid Leach Process 7-24 High Temperature Sulfuric Acid Leach Process - Services Profile 7-25 List of Tables (cont'd.) Number Table 7.3 Table 7.4 Table 7.5 Table 7.6 Table 7.7 Table 7.8 Table 8.1 Table 8.2 Table 8.3 Table 8.4 Table 8.5 Table 8.6 Table 8.7 Table 8.8 Table 9.1 Table 9.2 Table 9.3 Table 9.4 Page Constit,uents of Solid Effluent Streams - High Temperature Sulfuric Acid Process 7-26 Composition of Liquid Effluent Streams - High Temperature Sulfuric Acid Process 7-27 Composition of Gaseous Effluent Streams - High Temperature Sulfuric Acid Process 7-28 Summary Flows and Distribution of Potentially Toxic Elements - High Temperature Sulfuric Acid Leach Process 7-33 Summary of Waste Streams - High Temperature Sul- furic Acid Leach Process 7-34 Concentration of Toxic Elements in Waste Streams - High Temperature Sulfuric Acid Process 7-35 Plant Inputs and Outputs - Reduction/Hydrochloric Acid Leach Process 8-25 Reduction/Hydrochloric Acid Leach Process - Services Profile 8-27 Constituents of Solid Effluent Streams - Reduction/ Hydrochloric Acid Leach Process 8-28 Composition of Liquid Effluent Streams - Reduction/ Hydrochloric Acid Leach Process 8-29 Composition of Gaseous Effluent Streams - Reduction/ Hydrochloric Acid Leach Process 8-30 Summary Flows and Distribution of Potentially Toxic Elements - Reduction/Hydrochloric Acid Leach Pro- cess 8-36 Summary of Waste Streams - Reduction/Hydrochloric Acid Leach Process 8-37 Concentration of Toxic Elements in Waste Streams - Reduction/Hydrochloric Acid Leach Process 8-38 Plant Inputs and Outputs - Smelting Process 9-31 Smelting Process - Services Profile 9-33 Composition of Solid Effluent Streams - Smelting Process 9-34 Composition of Liquid Effluent Streams - Smelting Process 9-35 List of Tables (cont'd.) Number Page Table 9.5 Composition of Gaseous Effluent Streams - Smelting Process 9-36 Table 9.6 Summary Flows and Distribution of Potentially- Toxic Elements - Smelting Process 9-42 Table 9.7 Summary of Waste Streams - Smelting Process 9-43 Table 9.8 Concentration of Toxic Elements in Waste Streams - Smelting Process 9-44 Table 10.1 Major Equipment Requirements and Approximate Sizes for Shipboard Partial Processing Plant - Reduction/ Ammoniacal Leach Process 10-12 Table 10.2 Reduction/ Ammoniacal Leach Partial At-Sea Pro- cessing Pregnant Liquor Production 10-13 Table 10.3 Reduction/Ainmoniacal Leach Partial At-Sea Pro- cessing Basic Metal Carbonate Production 10-13 Table 10.4 Major Equipment Requirements and Approximate Sizes for Shipboard Partial Processing Plant High Tem- perature H2SO4 Leach Process 10-14 Table 10.5 High Temperature Sulfuric Acid Leach Partial At- Sea Processing Pregnant Liquor Production - Services Profile 10-15 Table 10.6 High Temperature Sulfuric Acid Leach Partial At-Sea Processing Mixed Metal Sulfide Production 10-15 Table 10.7 Major Equipment Requirements and Approximate Sizes for Shipboard Partial Processing Plant - Reduction/ HCl Leach Process 10-16 Table 10.8 Reduction/Hydrochloric Acid Leach Partial At-Sea Processing Pregnant Liquor Production - Services Profile 10-17 Table 10.9 Major Equipment Requirements and Approximate Sizes for Shipboard Partial Processing Plant - Cuprion/ Ammoniacal Leach Process LO-18 Table 10.10 Cuprion/Ammoniacal Leach Partial At-Sea Processing Pregnant Liquor Production 10-19 Table 10.11 Cuprion/Ammoniacal Leach Partial At-Sea Processing Basic Metal Carbonate Production - Services Profile 10-19 Table 10-12 Process Plant Inputs and Outputs At-Sea Partial Pro- cessing for Production of Pregnant Liquor 10-24 xi List of Tables (cont'd.) Number Page Table 10.13 Process Plant Inputs and Outputs At-Sea Partial Processing with Reduction of Metals from Preg- nant Liquors 10-26 Table 10.14 Weekly Tonnage for At-Sea Processing (short tons) Pregnant Liquor Production 10-30 Table 10.15 Summary of At-Sea Process Vessel and Transporta- tion Ships - Pregnant Liquor Production 10-31 Table 10.16 American Flag Ships Available 10-34 Table 10.17 Large U.S. Flag Tankers (over 51,000 DWT) 10-35 Table 10.18 Land Plant Inputs and Outputs - Reduction/ Ammoniacal Leach Process At-Sea Processing Pregnant Liquor Production 10-37 Table 10.19 Land Plant Inputs and Outputs - Reduction/ Ammoniacal Leach Process At-Sea Processing - Basic Metal Carbonate Production 10-38 Table 10-20 Land Plant Inputs and Outputs - High Temperature Sulfuric Acid Leach Process At-Sea Processing - Pregnant Liquor Production 10-40 Table 10-21 Land Plant Inputs and Outputs - High Temperature Sulfuric Acid Leach Process At-Sea Processing - Mixed Metal Sulfide Production 10-41 Table 10-22 Land Plant Inputs and Outputs - Reduction/Hydro- chloric Acid Leach Process At-Sea Processing - Pregnant Liquor Production 10-42 Table 10-23 Land Plant Inputs and Outputs - Cuprion/Ammoniacal Leach Process At-Sea Processing - Pregnant Liquor Production 10-44 Table 10-24 Land Plant Inputs and Outputs - Cuprion/Ammoniacal Leach Process At-Sea Processing - Basic Metal Carbonate Production 10-45 Table 11.1 Classification of Wastes by Chemical Characteris- tics - Nodules Processing 11-2 Table 11.2 Classification of Wastes by Physical Characteris- tics - Nodules Processing 11-3 Table 11.3 Materials and Utilities Summary - Waste Treatment, Reduction/Ammoniacal Leach Process 11-15 Xll List of Tables (cont'd.) Number Table 11.4 Table 11.5 Table 11.6 Table 11.7 Table 12.1 Materials and Utilities Summary - Waste Treat- ment, High Temperature Sulfuric Acid Leach Process Materials and Utilities Summary - Waste Treat- ment, Reduction/Hydrochloric Acid Leach Pro- cess Materials and Utilities Summary - Waste Treat- ment, Smelting Process Material and Utilities Summary - Waste Treat- ment, Cuprion/Ammoniacal Leach Process Estimated Noise Level at 1000 Feet from center of Processing Activity Page 11-16 11-17 11-18 11-19 12-7 I xiii List of Figures Number Page Figure 2.1 Process Alternatives 2-2 Figure 3.1 Capital Cost Estimates 4-4 Figure 3.2 Operating Cost Estimates 3-4 Figure 5.1 Process Outlines 5-2 Figure 5.2 Ore Preparation and Drying 5-4 Figure 5.3 Reduction 5-5 Figure 5.4 Leaching 5-7 Figure 5.5 Tailings Washing 5-8 Figure 5.6 Liquid Ion Exchange/Extraction 5-9 Figure 5.7 Liquid Ion Exchange/ Stripping 5-10 Figure 5.8 Copper Electrowinning 5-12 Figure 5.9 Nickel Electrowinning 5-13 Figure 5.10 Cobalt Recovery 5-14 Figure 5.11 Ammonia Recovery 5-16 Figure 5.12 Materials Handling 5-17 Figure 5.13 Services 5-18 Figure 5.14 Reduction/Ammoniacal Leach Process: Distribution of Potentially Toxic Constituents of Manganese Nodules 5-38 Figure 6.1 Process Outlines 6-2 Figure 6.2 Ore Preparation 6-5 Figure 6.3 Reduction/Leach 6-6 Figure 6.4 Oxidation Leaching 6-7 Figure 6.5 Wash-Leaching 6-8 Figure 6.6 Liquid Ion Exchange/Extraction 6-9 Figure 6.7 Liquid Ion Exchange/Stripping 6-11 Figure 6.8 Copper Electrowinning 6-12 Figure 6.9 Nickel Electrowinning 6-13 Figure 6.10 Cobalt Recovery 6-15 Figure 6.11 Ammonia Recovery 6-16 Figure 6.12 Materials Handling 6-18 XIV List of Figures (cont'd.) Number Figure 6.13 Figure 6.14 Figure Figure Figure Figure Figure Figure Figure Figure Figure Figure Figure Figure Figure Figure Figure Figure 7.1 7.2 7.3 7.4 7.5 7.6 7.7 7.8 7.9 7.10 7.11 7.12 7.13 7.14 7.15 7.16 Figure 8.1 Figure 8.2 Figure 8.3 Figure 8.4 Figure 8.5 Figure 8.6 Figure 8.7 Figure 8.8 Services Cuprion/Ammoniacal Leach Process: Distribution of Potentially Toxic Constituents of Manganese Nodules Process Outline Ore Preparation Leaching Tailings Washing pH Adjustment Copper LIX Extration Copper Electrowinning Cu Raffinate Neutralization Nickel LIX Extraction Nickel LIX Stripping Nickel Electrowinning Cobalt Recovery NH3 Recovery Materials Handling Services High Temperature Sulfuric Acid Leach Process: Distribution of Potentially Toxic Constituents of Manganese Nodules Process Outline Ore Preparation and Drying Hydro chlorinat ion Leaching /Washing Copper Extraction and Stripping Copper Electrowinning Cobalt Extraction Nickel Extraction Chloride Scrubbing and Stripping Page 6-19 6-28 7-2 7-5 7-6 7-7 7-8 7-9 7-11 7-12 7-13 7-15 7-16 7-17 7-19 7-20 7-21 7-?9 8-2 8-5 8-6 8-8 8-9 8-11 8-12 8-13 XV List of Figures (cont'd.) Numb er Page Figure 8.9 Nickel Electrowinning 8-15 Figure 8.10 Manganese Recovery ' 8-16 Figure 8.11 Cobalt Recovery 8-17 Figure 8.12 HCl Recovery 8-19 Figure 8.13 Waste Recovery 8-20 Figure 8.14 Materials Handling 8-22 Figure 8.15 Services 8-23 Figure 8.16 Reduction/Hydrochloric Acid Leach Process: Distribution of Potentially Toxic Constituents of Manganese Nodules 8-31 Figure 9.1 Process Outline 9-2 Figure 9.2 Ore Preparation and Drying 9-5 Figure 9.3 Reduction 9-6 Figure 9.4 Smelting 9-8 Figure 9.5 Converting 9-9 Figure 9.6 Ferromanganese Reduction 9-11 Figure 9.7 Leaching/Washing 9-12 Figure 9.8 pH Adjustment 9-13 Figure 9.9 Copper LIX Extraction/Stripping 9-14 Figure 9.10 Copper Electrowinning 9-16 Figure 9.11 Cu Raffinate Neutralization 9-17 Figure 9.12 Nickel LIX Extraction 9-18 Figure 9.13 Nickel LIX Stripping 9-19 Figure 9.14 Nickel Electrowinning 9-21 Figure 9.15 Cobalt Recovery 9-22 Figure 9.16 NH3 Recovery 9-24 Figure 9.17 Waste Treatment 9-25 Figure 9.18 Materials Handling 9-26 Figure 9.19 Services Figure 9.20 Smelting Process: Distribution of Potentially Toxic Constituents of Manganese Nodules 9-37 xvx List of Figures (cont'd.) Number Page Figure 10.1 At Sea Processing Plant with Hose to Mining Ship 10-A Figure 10.2 Process Outlines - At-Sea Preprocessing - Reduction/Ammoniacal Leach Process 10-8 Figure 10.3 Process Outline - At-Sea Preprocessing - High Temperature Sulfuric Acid Leach Process 10-9 Figure 10. A Process Outline - At-Sea Preprocessing - Reduction/Hydrochloric Acid Leach Process 10-10 Figure 10.5 Process Outlines - At-Sea Preprocessing - Cuprion/Ammoniacal Leach Process 10-11 Figure 11.1 Generalized Process Schematic - Waste Treatment 11-8 Figure 11.2 Waste Treatment - Reduction/Ammoniacal Leach Process 11-9 Figure 11.3 Waste Treatment - High Temperature Sulfuric Acid Leach Process 11-10 Figure 11.4 Waste Treatment - Reduction/Hydrochloric Acid Leach Process 11-11 Figure 11.5 Waste Treatment - Smelting Process 11-12 Figure 11.6 Waste Treatment - Cuprion/Ammoniacal Leach Process 11-13 Figure 12.1 Estimated Noise Levels of Equipment Contributing to Ambient Sound Reduction/Ammonia Leach Process 12-2 Figure 12.2 Estimated Noise Levels of Equipment Contributing to Ambient Sound High-Temperature Sulfuric Acid Leach Process 12-3 Figure 12.3 Estimated Noise Levels of Equipment Contributing to Ambient Sound Reduction/Hydrochloric Acid Process 12-4 Figure 12.4 Estimated Noise Levels of Equipment Contributing to Ambient Sound Smelting Process 12-5 Figure 12.5 Estimated Noise Levels of Equipment Contributing to Ambient Sound Cuprion/Ammoniacal Leach Process 12-6 Figure 12.6 Typical A-Weighted Sound Level As Measured by a Sound Level Meter 12-8 k XV 11 1.0 INTRODUCTION 1.1 GENERAL Manganese nodules are potato-shaped concretions found on the deep ocean floor in many parts of the world. Although somewhat variable in composition, these nodules may become an important source of nickel, copper, cobalt, and manganese. Manganese nodules could be regarded as a relatively high grade nickel-copper-cobalt oxide ore since these three metals are the most valuable recoverable constituents of the nodules. While deposits of manganese nodules occur in every major ocean, those in the near-equatorial North Pacific are of greatest interest at present because of the greater nodule density in many deposits and their higher than average concentrations of metal values. High concentrations of nodules have been observed over many thousands of square miles of ocean floor and have been estimated to average 10 kg (dry weight)/m2 (2 lb/ft2) in mineable areas. There is, therefore, an enormous potential reserve. The deposits of greatest commercial interest lie in a belt south of the Hawaiian Islands and north of the Equator and extend almost from Mexico to 180°W. Typically, these deposits lie at depths between 4,000 and 6,000 meters (13,000 and 20,000 feet). Because of economic and strategic incentives and U.S. leadership in the required technology, the United States will probably be the leader in the development of the manganese nodule deposits found in deep ocean basins. Because of proximity to our shores and high concentration of metal values, the deposits south of Hawaii appear to be the most likely targets, at least for first generation operations. In general, an ocean mining operation will consist of mining vessels, a fleet of carriers which will transport nodules (ore) to a port facility and an on-shore processing plant. There is a possi- bility that at-sea (shipboard) processing or partial processing may be feasible in the future. However, the requirement to develop new technology for at-sea processing will probably lead to use of onshore plants in first generation systems. Since initial nodule mining operations will probably be centered in the belt south of the Hawaiian Islands, the West Coast, Hawaii, or possibly the Gulf Coast may be considered by industry for the siting of port facilities and onshore processing plants. The industry's decisions will be based on total system economics and regulatory considerations in potential areas. 1.2 NOAA ENVIRONMENTAL PROGRAM It is likely that one or more Federal actions will be associated with development of a deep ocean mining industry in the United States and would be considered a "major Federal action" under the National Environmental Policy ACT (NEPA) . Thus, it is likely that the Federal Government will be required to prepare one or more environmental impact statements for such an industry to develop. As a result, the U.S. Department of Commerce's National Oceanic and Atmospheric Administration (NOAA) lias undertaken two studies to 1-1 assess the potential impacts of the new industry. The Deep Ocean Mining Environmental Studies (DOMES) Project, already underway, is designed to assess the potential environmental effects of at-sea nodule mining operations in the Pacific area of commercial interest and to develop information for the development of appropriate environmental safeguards. In a complementary pro- ject, the NOAA Office of Marine Minerals has initiated a three-phased program to assess the environmental and socio-economic impacts of other activities associated with development of the industry, specifically, ore (nodule) trans- portation, processing of nodules and disposal of process wastes. Review of published information dealing with the broad topic of manganese nodules reveals that little or no work has been done, or at least divulged to the general public, on the potential environmental impacts associated with the transportation and processing of manganese nodules. Since this may well become a major industry, both positive and negative environmental effects may be significant. Failure to understand, anticipate, and control or mitigate potential adverse impacts could also lead to delays and add to the total cost of development of the industry. Definitions of these impacts at an early stage can also aid industry by permitting the design of operations which are environ- mentally compatible and economically attractive to the maximum practical extent. The processing of manganese nodules has the following unique aspects of impor- tance; • The processing will contain some elements of existing technology but, as a whole, represents a new type of industry. • The operations, as a class, contain elements of existing technology known and used throughout the nonferrous metals industry, but such operations have in general not been conducted in the coastal states where manganese nodule processing sites would probably be sought. • While it appears, a priori, that nodule processing is not unusually hazardous and does not present unknown dangers, it is unusual for a new industry to develop in full public view. The objective of Phase I of the NOAA processing impact assessment is to describe and characterize the relevant activities associated with at-sea and land transportation, processing, and waste disposal in a manner suitable for use in subsequent environmental and socio-economic impact studies. The Phase I work was performed under two separate, competitively awarded contracts. The work on processing systems was designated "Phase lA" and the work on transpor- tion and waste disposal was designated "Phase IB". Contract work on Phase lA began during July 1976. Ultimately the NOAA processing study will result in the assessment of potential environmental and socio-economic impacts of locating processing plants and associated facilities in U.S. geographical areas whose charac- teristics are representative of those of areas in which the industry may seek to locate. 1-2 1.3 SCOPE OF THIS STUDY In order to permit the planned assessment it was necessary to use infor- mation from relevant literature, supplemented by the considerable use of analogies and engineering judgment, to answer the following questions for Phase lA: • What process routes are likely to be employed? Obviously unattrac- tive routes can be easily elminated, but it will not be possible to identify a best or optimum route or define any given route with a high degree of detail. • At what scale of operation will the first generation processing plants operate? Given the lack of definitive cost estimates published, it will be possible to define likely capacities with an uncertainty of the order of ± 33%. • What are likely plant inputs and outputs? Comprehensive material and energy balances can define inputs and outputs within an accuracy of ±10-20%. • What supporting services will be required? Definition of the process route and material and energy balances along with the characterization of major Items of equipment and outputs will permit estimates to be made of support services such as transport, infrastructure, and waste containment requirements based on analogies to similar operations. • What is the amount and nature of the waste products produced including the distribution of potentially toxic or hazardous substances? The amounts of waste will have been defined in the material balance to within ± 10%, and the distribution of potentially toxic elements can be deter- mined, although w*ith a somewhat higher degree of uncertainty in the absence of experimental data. These estimates will permit a preliminary assess- ment of the hazards of nodules processing relative to other well known industrial operations. • What are the potential hazards involved from an accident in an operating nodules plant? Well established techniques can be used to identify the hazards of accidental release and their probability. • How will wastes be treated to facilitate their safe disposal? Having defined, qualitatively, the nature of the waste and knowing the amounts involved, preliminary assessments can be made of process possibilities which will render them safe for disposal. • How will at-sea processing or partial processing of nodules affect the charac- teristics of a land based plant? Given estimates of the state of technology in first generation plants and what may be done on shipboard, it is possible to examine the potential for splitting process operations between land and sea and determining the impact of the various alternatives. 1-3 These questions, and others, are addressed in this volume. The results are summarized and interpreted in Volume I to suit a general readership and considerable detail has been eliminated. In this volume, however, the results are presented in detail with little attempt to bring perspective to the data presented. It is anticipated that this volume will be of most interest to the technical reader who will be able to bring the necessary perspective from his own experience. The objective of retaining this detail is to permit the reader to have access to the documentation of all assumptions which were made in defining the process routes and developing the detailed material and energy balances and process characterizations. In order to help assess the "realism" and "relevancy" of the work, NOAA. requested industry, environmental interest groups, and interested agencies (e.g., the Department of the Interior's Bureau of Mines and the U.S. Environ- mental Protection Agency) to review the work and to constructively comment on its results. This was done through a series of public briefings which were attended by industry, government, and outside processing experts as well as by representatives of environmental interest groups and the public. In the use of the information developed during Phase lA it should be recognized that : • This is a "paper" study and the exact nature of the processes which will be used by each consortium, and their environmental and other implications, will not be known until details are released by industry in context of applying for permits or authorizations to engage in commercial operations under a yet to be defined international or domestic legal regime. • This work is designed to serve as an input for subsequent impact studies and not to assess potential impacts themselves; however, it is important to recognize the need to consider the reported results in the overall context of other industries and activities which involve similar operations and use of similar materials. 1-4 I 2.0 ALTERNATIVES FOR ON-SHORE PROCESSING OF MANGANESE NODULES 2.1 BASIS FOR THIS STUDY The economic potential for metal recovery from deep sea nodules has long been recognized, but the technology needed to engage in commercial scale mining operations has yet to be fully demonstrated. It is beyond the scope of this work to undertake either a complete economic evaluation of nodules processing Alternatives or to define the "optimum" mining or processing system. In all likelihood no optimum exists: rather, special circumstances will favor the adoption of one or another route to suit the re- quirements of a particular developer. Thus, we will treat the economics of nodules processing in a general way, drawing heavily on analogies to the economics of similar operations insofar as possible and using accepted, though approximate, techniques to extrapolate and interpolate results. Specifically, the economics of nickel recovery from lateritic ores will be taken as a reference point for capital and operating cost estimates. It is assumed, then, that nodules processing must show some advantage vis-a-vis laterites processing to justify the risks attending the development of new technology, particularly mining in this case. No process alternative will be rejected from further consideration unless it can be shown clearly that some inherent characteristic renders it specific to a requirement unlikely to be met, e.g., the availability of reductant/lixiviant/ energy at higher rates or lower costs than probable in a limited geographic area. While all process alternatives will be considered, detailed evaluations (Task II) will require that some design basis be available; it is beyond the scope of this study to invent new processes. Thus, only those processes for which at least some data on metal recovery schemes, efficiencies, and required process conditions have been revealed can be evaluated. Even so, it is anticipated that information on practices and results of related technology, e.g., laterites, will be a key source of design criteria. The time frame hypothesized for operation of the plants described in this study is the 1980 's and it will be assumed that all pilot scale work will have been completed. This would include demonstration of the mining system at a scale sufficient to guarantee an assured supply of nodules at the rate required by the commercial plant. It might also have included construction and operation of a relatively large pilot plant (or relatively small commercial plant) in the general area chosen for the commercial facility. We will assume, however, that the processes evaluated will be reflected in plants which are both new and self-sustaining. Finally, this evaluation will be restricted to the requirements and character- istics of the process plant itself. An assured supply of nodules will be assumed as well as the availability of a site suitable for disposal of process wastes, the amount and nature of which will be defined by this study. 2.2 PROCESSING ALTERNATIVES 2.2.1 Classification Schemes A large number of potential processes have been studied in an attempt to define an economic route for recovering metals from nodules. Generally these processes must include a step to reduce manganese from the +4 to +2 oxidation 2-1 state in order to break down the matrix of the nodules and allow the other metal oxides to be extracted by suitable means. Meyer-Galow suggests a number of processing alternatives in the wheel shown in Figure 2.1. We have chosen to classify the processes either according to the lixiviant used in the roast or leach, or as a smelting operation. The lixiviants fall in three classes, those producing acid sulfate solutions, chloride (halide) solutions and ammonia/ammonium solutions. Any process using ammonia /ammonium leach must also include a reducing step since this lixiviant does not act as a reducing agent. In addition, any process may incorporate additional reducing agents or oxidizing agents in an effort to separate the metal products selectively. The bibliography on processes includes a cross-reference section which reflects this classification. In addition, the cross-reference section includes subsections on metal separation, overview articles, process articles which do not fit other classifications, and articles by major groups or organizations who have published results on the processing of manganese nodules i 2.2.2 The Influence of External Variables By external variables, we mean factors beyond those directly associated with the extractive metallurgy of nodule processing. The most obvious external variables are the grade and reserves of the nodule ores available. While the total resource appears to be vast, it is less clear that there exist large reserves of nodules of a composition which uniquely favor one processing route above all others. Some nodule compositions have been reported, and their ore values at current prices, are summarized in Table 2.1. Both the total value and the distribution vary widely. Thus a nodule of the mid Pacific (10°S, 147°W) type, which contains a high nickel/copper/cobalt value, would favor a so-called three metal process if available in sufficient quantity. Similarly, a nodule of the "southeast" Pacific type would depend strongly on manganese recovery for profitability. FIGURE 2.1. PROCESS ALTERNATIVES (FROM MEYER-GALOW ET AL . (79) 2-2 The Atlantic type, high in acid -consuming basic materials would be most economically processed using a basic lixiviant, or perhaps by smelting, so as to minimize consumption of process materials. Other than ore grade, perhaps the potentially most important external variable would be the existence of a special situation with regard to either the supply of a reductant or lixiviant or for the disposal of a by-product. By special situation we mean the existence of a business arrangement, specific to the process and size under consideration, which overcomes a serious technical or economic problem. A substantial credit received for the values contained in solubles purge streams, which would otherwise be disposal problems, or the ability to accept others' waste or by-product streams for use as reductants/lixiviants at little or no cost would be examples of these situations. It should be noted that, if these do develop, they will undoubtedly dictate plant sitings. It is to be expected that nodules processing plants will be relatively large and will be located conveniently to raw materials and consumables, including sources and disposal facility for process wastes. Local real estate and/or support capabilities may dictate either the scale or processing technique suitable for a given site. Finally, the existence of special marketing situations could dictate the process route chosen and/or the specifications for the products produced. Marketing considerations, including the need to introduce new product forms and an assurance of their acceptance, are most likely to bear on the decision as to the attractiveness of manganese recovery. This, in turn, would dictate the plant size because of the impact of the market share relationships between manganese and all other metals. TABLE 2.1 Some Nodules Compositions and th eir Va lues as res at Current Prices^ ^ Bl Atlantic ake Plateau Pacific West 10' S, 147°W "Southeast" Avg. of 54 Indian % Manganese 17 25.6 19.81 24.2 13.56 % Nickel 0.6 1.5 0.96 0.99 0.32 % Copper 0.1 1.2 0.31 0.53 0.1 % Cobalt 0.3 0.18 0.16 0.35 0.36 % Molybdenum N.R. N.R. 0.04 0.05 0.03 % Iron 11 5.0 10.2 14 15.75 % Silica 1.7 N.R. N.R. 9.4 N.R. % Calcium 13 N.R. N.R. 1.9 N.R. Value, $/T, of Contained Manganese 85.0 128.0 99.1 121.0 67.8 Nickel 24.0 60.0 38.4 39.6 12.8 Cobalt 24.0 14.4 12.8 29,0 28.8 Copper 1.5 18.0 4.7 8.0 1.5 Molybdenum , $/T — — 1.5 156.5 1.9 198.5 1.1 Total 134.5 220.4 112.0 Ni/Cu/Co Total, $/T 49.5 92.4 55.9 75.6 43.1 Compositions from various sources (20) (29) (36). ^^Vid 1976. Mn @ $.25/lb, Ni @ $2.00/lb, Cu (3 $.75/lb, Co (? $4.00/lb. Mo @ $1.85/lb. 2-3 2.2.3 Acid Sulfate Lixiviant Processes Processes for which acid sulfate is the lixiviant include low and high temperature sulfuric acid leaches of raw nodules, leaching of nodules which have been subjected to sulfation roasting or acid baking, and direct reduction/ leaching employing sulfur dioxide. Low temperatures , 24-100°C, sulfuric acid processes offer the potential of relatively mild leach conditions and have been extensively studied. Brooke (18) reports leaching experiments in which basic constituents consumed ten times the acid required to solubilize the metals, and recoveries no higher than 65%. Fuersteneau (37) and Han (44) report up to 90% recovery of copper and 60% for nickel in leaches of 48 hours duration on 1 micron particles at pH 1 . Han also found (42) up to 74% nickel, 80% copper recoveries after 7 days on 100 mesh solids with the consumption of 0.3 lb acid/lb nodules. Kane (59) (67) (68) reports the use of staged leaches involving sulfuric acid followed by ferrous sulfate as a reductant for manganic oxide. Overall, greater than 80% recoveries of all metals - including manganese - are obtained at the expense of large consumptions of ferrous sulfate. In the absence of a reductant, low temperature sulfuric acid leaches do not solubilize appreciable manganese. The situation in high temperature leaching is less clear. Brooks (20) reports "good" recoveries of copper/nickel/cobalt but "poor" manganese recovery at 200°C in the presence of oxygen. Hanig (46) and Neuschutz (61) report excellent recoveries of copper and nickel, 80-90%, and good cobalt recoveries, 70-80% at 200-250°C in the absence of oxygen with acid consumptions of about .3 lb/lb nodules. Ulrich's recoveries of cobalt and acid consumptions were somewhat lower, 40% and .13 lb/lb respectively (125). In all these studies iron solubilizations were low, 1-2%^ and of the order of 5% of the nodules' manganese was leached. Han (45) presents data which show that a high oxygen pressure, 100 psi, is required to minimize manganese and iron solubilization and keep acid consumption to .03 lb/lb nodules. At lower oxygen pressures, ca. 50 psi, 15% of the manganese dissolved, and acid consumption reached 0.3 lb/lb. While these processing conditions are severe, they are known commercially in laterites processing at Moa Bay, and recoveries and selectivities , as shown, can be excellent. Sulfur dioxide gas can serve, like ferrous sulfate, as a reductant for the manganic oxide and as a lixiviant in leaching. Brooks (20) obtained in excess of 97% solubilization of manganese/nickel/copper/cobalt and 72% iron dissolution by bubbling 0.8 lb S02/lb nodules into a leach slurry. Kane's patent (64) shows similarly high solubilization of these metals. Kane (61) also reports that sulfation roasting with sulfur dioxide followed by staged leaching, with and without oxygen, can selectively leach manganese and then nickel/copper/cobalt. Pure sulfur dioxide or synthetic roaster gas are effective reductants and permit solubilization of greater than 90% of the contained nickel/cobalt/ manganese. High copper recovery, however, required subsequent acid leaching. Van Heck (126) sorbed sulfur dioxide at low concentrations (simulating stack gases) onto nodules at 300-500°C and water leached the residues. High recoveries were obtained for all metals except iron with Atlantic nodules, but recoveries were lower, only 45% nickel, for Pacific nodules. Brooks (13) also roasted nodules at 400-725°C with 10-20% SO2 gases. The solubilizations of manganese and cobalt were very high ( 93%) but nickel (73%) and copper (77%) were lower and sulfur dioxide consumption was still high. Kane (42) and Meyer-Galow (58) show that pyrrhotite or pyrite are effective sulfating agents between 350-650°C. Near complete solubili- zations of manganese and cobalt and high ( 80%) solubilization of nickel and copper are possible. 2-4 1.1. k Chloride (Halide) Lixiviant Processes The use of a halide as a lixiviant usually produces the partial dissolution of all metal components in nodules including manganese. As such, these processes should be considered separately from the other lixiviants, since their economics will depend on the recovery and sale of manganese as well as nickel, copper and cobalt. Halides, as a broad category, are discussed in a number of citations as possible lixiviants but for practical purposes this invariably means chlorides, and little or no data are available for any other halide. Processes for which chlorides are used as a lixiviant include gaseous HCl leach, halidation of nodules in molten salt bath and chloride segregation roasting. Kane and Cardwell in a number of patents and a paper (25,29,57,66) proposed a number of methods for a HCl roast. Cardwell (25) described a process based on HCl roast which recovers 99.7% Cu, 97% Co, 99.9% Mn and 96.7% Ni and is under study in pilot plant testing for commercialization. Cardwell (29) and Kane (57) described a process where nodules are roasted at 500°C with HCl vapor for 2 hours, cooled to 300°C, blown with air saturated with water for 30 minutes and leached for 1 hour with aqueous solution at pH 2 at a weight ratio of nodules to leach liquid of 1 to 1.3. The recovery was 95% Mn, 97.4% Ni, 96% Cu , 97.6% Co and no Fe . Cardwell also teaches that optimum temperature range is 500-600°C. Kane (66) reports that as the temperature is increased above 500°C to 1000°C, all the metal chloride constitutents are vaporized completely and swept out by HCI/CI2 flow. The CI2 vapor is produced largely by the reduction of Mn . Cardwell (29) suggested that the vaporization can take place in temperature stages causing fractional separation of the chlorides. In one experiment, the vapor was separated into three streams based on temperature: Temperature Range % of Recovered Metal Mn Ni Cu Co Fe Below 550°C 5 20 70 20 92 550-850°C 65 80 30 80 8 850-1000°C 30 Thus, the high temperature stream contains essentially pure manganese chloride and no other metal. In general, the aqueous HCl leaching gave less recovery than the HCl roast. Ulrich (125) reports some intermediate temperature (100-200°C) aqueous HCl leaches. His best results show a recovery of 85-90% Cu, 85-98% Ni, 27% Fe, 40-60% Co and 29-40% Mn at 200°C in the range of 0.2 to 0.4 HCl-to-nodule weight ratios. Higher concentrations of HCl dissolved larger portions of the manganese. Cardwell (27) mixed NaCl and H2SO4 to generate HCl in-situ which reacted with nodules in several experiments with temperatures ranging from 300-525°C for one hour. The mixture was then leached with water. The results showed a large portion of manganese (67-93%) , copper (84-95%) and cobalt (83-88%) recovered and lesser amounts of nickel (71-83) and iron (73-0) . The iron chloride apparently evaporated at 525°C. Heating similar mixtures to 950°C for 2 hours, with coal and chlorine gas added, dissolved greater than 98% of the Fe , Cu, Ni and Co and 86% of the Mn. Similar results were reported by Kane (40,48) showing that complete dissolution does not occur below 900°C and under 2 hours. 2-5 Kane (65) teaches a process to convert metals in nodules to chlorides in a molten salt bath made up of alkali metal salts and/or alkaline earth salts. The contact is made for 6 hours at 600-900°C followed by an aqueous leach. The recoveries of metals were 80% I'ln, 76-85% Ni, 92-94% Cu, 72-80% Co and no Fe present. These results do not compare favorably with the gaseous HCl roast for the same conditions. However, the process does prevent the dissolution of iron. Hoover (49) suggested a segregation process. where the nodules are heated with various metal chlorides and reducing agent (coke) between 700-1000°C. CaCl2 gave the best results at 850°C with relatively poor recoveries: 75% Cu, 25% Ni and 25% Co. 2.2.5 Ammoniacal Lixiviant Processes Processes for which ammonia/ ammonium compounds are the lixiviant require the nodules to be reduced prior to leaching in order to obtain acceptable recoveries. Reduction may be carried out using either pyro or hydrometallurgical techniques. Significant efforts have been directed towards studying reduction/leaching procedures which are similar to those employed in the Caron process for recovering nickel from laterites as practiced at Nicaro. Bare (9), Brooks (19), Han (43), and Redman (93) have roasted nodules in carbon monoxide/dioxide mixtures ranging in composition from 2:5 to 2:1 as well as in carbon monoxide/hydrogen and simulated syngas mixtures at temperatures between 350°C and 800°C. Reduction times ranged from about 1 to 6 hours with nodules ground to between 20 and 100 mesh. The calcines were generally cooled in a reducing atmosphere following roasting, and aeration was sometimes used during leaching. It has been found that solubilizations of nickel/copper /cobalt ( and molybdenum) are generally good, in the range of 80-90% for copper and nickel and 50-70% for cobalt (70-80% for molybdenum) , at low temperatures (R.T.to 65°C) in solutions containing of the order of 100 g/1 NHo and 30-80 g/1 CO2. The rate of copper dissolution is quite rapid, with typically 80% being obtained within one hour's leach, although higher roasting temperatures decrease extractions. Increasing leach temperatures significantly increases the rate of nickel solubilization as does an increase in roasting temperature up to about 750°C. Acceptable solubilizations were usually obtained during the course of a few hours leaching, but much longer times (32 hours) did increase solubilization, particularly of nickel and cobalt. It was found that suspension of the calcine in tbe leach solutions resulted in the initial solubilization of some manganese and iron which subsequently reprecipitated . Also, some manganese carbonate was precipitated by the sorption of carbon dioxide from solution by manganous oxide. High ammonia/carbon dioxide concentrations result in the solubilization of significant (to 80%) amounts of manganese. Han (43) and Rolf (95) also show that manganese can be solubilized, to the exclusion of nickel and cobalt, if desired, by leaching in ammonium sulfate ( as opposed to ammonium carbonate) following reduction with either carbon monoxide/ dioxide mixtures or hydrogen. Wilder (132) (133) showed that roasting a mixture of coal (3-6%) and nodules at 650-800°C is an effective reducing procedure. Leaching the calcine in ammonium carbonate with air sparging at room temperatures for 4 hours gave 85-90% solubili- zations of all desired metals, including molybdenum. Skarbo (101) (102) showed that metals can be solubilized in ammonia/ammonium systems from nodules which have been reduced hydrometallurgically . Manganous sulfate is an effective reductant at 60°C, taking manganese from +4 to +3, and solubilizing 2-6 greater than 87% of the copper plus nickel. Alternately, in leaches at 200°C, ammonium chloride, sulfate or carbonate solubilized 70-90% of the copper plus nickel. Maintaining a carbon monoxide pressure during the leach precipitated manganese carbonate. Szabo (120) and Pemsler (90) described a hydrometallurgical reduction in which the cuprous ion reduces manganese, and is in turn re-reduced by carbon monoxide, the net effect being formation of manganese carbonate. In a staged reduction/leaching system, 97% reduction was obtained with greater than 92% solubilization of copper and nickel. While reduction/leaching conditions were mild (40-70°C, 1-2 hours, 100 g/1 NH3, 80 g/1 CO2) , slurry densities were quite low. 2.2.6 Smelting Smelting processes require high temperatures (1200°C) as well as the use of a reductant either prior to and/or during the smelting operation. Vasilchikov (127), Beck and Messner (13) and Sridhar (114) all reported smelting tests. In these tests, the nodules were heated in crucibles for 0.3-2 hours, and reductant (either graphite, coke or carbon) was added. In Sridhar' s tests (78) , a reducing atmosphere simulating the composition of a combusted fuel oil was also used. The results of these tests are summarized below: % Recovery of Metals Temperature (°C) Ni Cu Co Mo Fe Mn Source 1140 43.8 41 1 43 2 N.R. 33 1.3 Sridhar (114) 1215 64.1 64 64 3 N.R. 60 2 5.7 Sridhar (114) 1425 98.4 81 6 93 2 93.0 22 6 0.16 Beck et al . (13) 1450 88.5 86 7 45 3 88.0 38 2 0.25 Beck et al. (13) 1600 - 1650 100.0 86 93 N.R. 97 8.7 Vasilchikov (127) The results show good recovery of nickel and copper at temperatures above 1425°C. Better recovery of cobalt can be achieved by raising the temperature but at the expense of increasing the recovery of iron and, to a lesser extent, the recovery of manganese. Both Beck and Sridhar experimented with adding 0-15% sulfur (as elemental sulfur, pyrite and pyrrotite) . The addition of sulfur increases the recovery of iron. In all tests, the alloy concentrate resulting from the smelting represented only a small percentage (9-20%) of the original nodules. Beck also performed three pilot plant tests using 500 lbs of air-dried nodules. The smelting time was six hours. The results are listed in the table above at 1425°C and are in close agreement with the laboratory results. In a later work, Sridhar (116) reported similar results. A selective reduction for one hour at 1000°C with a reducing gas (simulating a rotary kiln with Bunker C fuel) was fallowed by a smelting at 1400°C. The recoveries were up to 98% Ni, 95%Cu, 98% Co, 90% Fe and 2.5%Mn. Sridhar also reported the recovery of minor constituents as: 86% Mo, 36% P, 64% As, 100% Sb, 0.5% V, 0.5% Ti, 6% Zn, 2% Pb and 1% Na . He also suggested that the manganese available in the slag could be converted into a commercial product. Sridhar dissolved the alloy by sulfiding and leaching with 100 g/1 H2SO4 at 110°C under a 150 psia oxygen pressure. Greater than 99% of the copper, nickel, cobalt and iron together with 80% of the minor constitutents were dissolved in 2 hours. Metal separation could follow using conventional processing, such as solvent extraction. 2-7 Wilder (134) teaches a smelting method in which Bunker C fuel, as a reductant, is added directly to the nodule charge without drying and ore preparation. The liquid reductant is apparently absorbed by the nodule. Using 3.5 wt% Bunker C fuel, the smelting was done at 1250-1350°C for 1 to 2 hours after a 4 hour heat-up time. Recoveries of 92.7% Cu, 99% Ni, 93.5% Co, 92% Mo, 37.5% Fe, and 0.047% Mn were obtained. 2.2.7 Miscellaneous Processes Limited information is available on attempts to recover metal values from nodules by methods not classified above. Halbach (40) proposes the use of sewage sludge as a reductant. While limited testing showed 75-85% metallization, the pelletized mixture had to be coated to prevent loss of the reductant. Hoover (49) attempted a segregation roast, mixing 10% each of coke and calcium chloride with the nodules, and roasting at 800-1050°C. Metal recoveries were poor, ranging from 20-70%. 2.2.8 Metals Separation and Recovery As discussed above, of the major nodules processing routes, only the HCl treatment requires recovering a manganese product. The reduction/ammonia leach process mobilizes nodule matrix manganese, but the resulting manganous Mn is rejected as carbonate or some form of oxide. The pyrometallurgical route yields a manganese-rich slag from which manganese can be recovered, or which can simply be held for stockpiling or disposal. In addition, all of the likely candidates for first-generation nodules processing plants reject iron, so that one is left, essentially, with the separation and recovery of the three principal value metals: nickel, copper and cobalt. Until satisfactory recov- eries of these metals is assured, it is probably premature, from the standpoint of process economics, to be concerned with the recovery of lesser metal values such as molybdenum and zinc. All of the major processing schemes lead to a solution containing the principal metal values, so that separation and recoveries are effected hydro- metallurgically . A review of published results indicates that nickel and copper recoveries are generally high, but that cobalt recoveries are low for the most part. This behavior may have influenced the inferred business philosophy that economic success hinges on high overall recoveries of nickel and copper and that substantial cobalt recovery is advantageous, but not essential. In most cases, it will be necessary to recover and recycle reagents. The cost of chemical reagents and their recovery is an important ingredient in the overall economics. The optimum metals separation and recovery scheme for a given initial metallurgical treatment of the nodules is influenced by the composi- tion of the solution produced. Thus, one may be dealing with an ammoniacal carbonate solution, a chloride solution or a predominantly sulfate solution. It is significant from a process standpoint to note that appreciable chloride will always be present unless pains are taken to comminute and thoroughly wash the nodules, which are'known to have substantial microporosity . Thus, 30-40% by weight of sea water in whole nodules means of the order of 1% of occluded sea salt. In some cases, system purge requirements may be determined by chloride buildup. In acid solution particularly, the introduction of occluded sea salt will be an adverse factor in corrosion of equipment. 2-8 2.2.8.1 Ammoniacal Solutions In a recent presentation (2) nickel and cobalt recovery from the ammoniacal solution produced by a fully hydrometallurgical reduction-leach process is described. It is indicated therein that sulfide precipitation had been investi- gated and judged to be uneconomical. The reason given was that this route leads to ammonium sulfate as an end-product, which would require the nodule processor to enter the fertilizer market or incur expensive reagent recycling costs. Similarly, a separation scheme based on carbonate precipitation was rendered unfavorable by excessive energy requirements, especially as steam for ammonia stripping. Several recent publications (2, 3, 100) outline a process involving sequential copper and nickel recovery by liquid ion exchange followed by electro- winning of the separate metals. No details are given of the cobalt recovery method, except for the suggestion that it can be removed chemically from the nickel-barren raffinate. One infers that the cobalt recovery method is not at as advanced a state of development as the nickel and copper recovery methods. Part of the generally experienced difficulty with the cobalt recovery may be due to inability to mobilize cobalt in the initial selective leach of raw nodules. Thus, the steady-state profile for a pilot plant operation (2) shows 6.2 g/1 Ni, 5.7 g/1 Cu and 0.2 g/1 Co in the pregnant leach liquor. This approximately 30 to 1 ratio of Ni to Co is to be compared with the ratio of occurrence in the nodules at approximately 6 to 1. A reducing-roast/ammonium carbonate leach has been applied to manganese nodules from the Atlantic (20). The recovery technique employed involved separation of copper by selective precipitation as the sulfide and separation of the nickel and cobalt as a mixed carbonate by stripping of NH^ and C02. It will be noted that the economic feasibility of each of these steps has been criticized . Skarbo and co-workers have patented a number of recovery methods applicable to separation of copper, nickel and cobalt from ammonical nodules leach liquors or mixed carbonates produced therefrom. In the latter category are: separation of nickel as the relatively insoluble ammine chloride following dissolution of the basic Cu, Ni, Co carbonate mixture, reduction to metal of part of the carbonate solids by hydrogen, HCl leaching of the remainder of the carbonate solids and reaction of the metal produced and the HCl leach solution with sulfur (104) . These processes have apparently been superceded by liquid ion exchange separation (106, 111, 112), including methods for stripping cobalt from oxime solution (105, 110). Finally, for ammoniacal systems, a variant of the reduction smelting process dissolves the Ni-Cu-Co metal in ammonia liquor (8) . Copper and cobalt are solvent- extracted from the leach solution, apparently separately, but by unspecified process steps. Nickel is recovered in the final step, which simultaneously regen- erates the ammonia. 2.2.8.2 Acidic Sulfate Solutions Separation and recovery possibilities for sulfate leach solutions containing Mn, Cu, Ni, Co and Fe and other impurities are succinctly summarized in Figure 2.1, taken from Meyer-Galow, e_t al. , (79) , who cite the following drawbacks of other methods : 2-9 • For cases where appreciable manganese is solubilized, as by a reduction leach with S02or Fe2+,its removal as MnlV hydrated oxide with CI2 (or air) and alkali incurs high cost for chlorine or high energy requirements for air oxidation Copper recovery by cementation (by use of Fe, Zn or Mn , for example) leads to an impure precipitate copper which must be refined for market. Extraction of copper by liquid ion exchange has various difficulties, including those of pH regulation and neutralization of the liberated H2SO4. Cobalt/nickel separation by chemical means is incomplete or uneconomical. Precipitation of cobalt by CI2 and alkali incurs coprecipitation of nickel. • Cobalt precipitation as carbonate is too costly, except where small amounts are involved. The solid ion-exchange approach to metals separation employs a Lewatit TP 207 resin (52). Its loading capacity is stated to be 50-55 kg metal per m^ of resin (70) . Manganese is the most weakly bound to this cation-exchanger so that on "filtration" of a pregnant leach liquor containing Mn (73 g/1) , Cu(2), Ni(3), Co(l), Zn ( 0.5), Fe ( 0.8), Al ( 0.05) as sulfates (pH 3.5) through the resin, the "filtrate" contains principally Mn2+, with only 2-5 ppm concentrations of Cu, Ni and Co. Separation of the metals retained by the resin is effected in four steps: Step 1: Copper (the most tightly bound) is eluted to 60 g/1 and electrowon. Step 2: Iron and aluminum are precipitated with Ca(0H)2 plus a small amount of CI2 to oxidize Fe^"*"; residual Mn"^ is also oxidized and precipi- tated in this step. Step 3: The Co + Ni + Zn filtrate from Step 2 is made strong in chloride, permitting zinc to be selectively extracted by tributyl-phosphate. Step 4: Cobalt is solvent extracted with triisononylamine, and nickel is precipitated from the raffinate. A differential, staged leach process (8), employing dilute sulfuric acid, purports to effect an initial preferential extraction of copper (and cobalt) from whole or coarsely crushed nodules. In the second state of dilute acid leaching, following a liquid/solid separation, nickel is said to be preferentially extracted. Mild leaching conditions, however, in the absence of a reducing agent to break down the nodules matrix, have not generally resulted in economically interesting percentage recoveries of the value metals (see 2,2.3). Two of the major contenders in the nodules processing arena yield sulfate leach solutions. This is obviously true for the high pressure leach process as described, for example, by Sisselman (100) and by Agarwal et_ a]^. (3). The separation scheme involves a liquid ion exchange/electrowinning sequence. More specifically, the pregnant leach liquor containing Cu, Ni, Co and 5 g/1 H2SO/ is neutralized with NH^ to pH 2, to permit selective separation of copper by means of LIX reagent. The raffinate from copper extraction is reported to be neutralized to pH 6 with NH3 , whereby it becomes feasible to extract the cobalt and nickel with LIX reagent. The two metals are selectively stripped from the organic phase and recovered by electrowinning from the respective strip solutions. The earlier mentioned drawback of converting sulfate to ammonium sulfate, and thereby consuming NH3 as well as H2SO4, applies to this process. 2-10 As the second example, the "pyrometallurgical" route, involving reduction smelting and sulfiding yields a Cu-Ni-Co-Fe matte which can be oxygen presssure leached in sulfuric acid (116) . Removal of iron from the leach solution is required and is effected by neutralization to produce an iron/gypsum precipitate. Copper is then separated and recovered by liquid ion exchange and electrowinning . The precise separation and recovery methods for nickel and cobalt have not been revealed. Sridhar e_t a]^. indicate only that cobalt "could be separated from nickel by solvent extraction or by selective oxidation and precipitation with base." They also suggest that cobalt and nickel "could be recovered as compounds or as metals by any of the several well-known methods." 2.2.8,3 Chloride Solutions The front end of the HCl metallurgical treatment of manganese nodules admits of several variations, including aqueous hydrochloric acid leach and hydrochlorin- ation at higher temperatures (29). These variants have in common the reduction of the manganese oxide matrix and the consequent mobilization of both major nodules metals, Mn and Fe , as their chlorides. The downstream extractive metallurgy is complicated by the need to separate out substantially all of the manganese and iron and to do this without detriment to the recovery of Ni, Cu and Co. To some degree, this feature of the HCl processes is common to all the reducing leach processes. The chloride route does permit certain specific separations not possible in the sulfate system: for example, extraction of FeCl3 by a water-immiscible organic solvent, such as a trialkylphosphate or an aliphatic amine (29), or separation of cobalt from nickel by selective solvent extraction of its chlorocomplex anion (70). However, the main feature that a relatively expensive reagent is utilized in solubilizing the nodules metals content in an essentially nonselective manner may offset specific advantages in metals separation. The variations which have been outlined in the patent and other literature on the hydrochlorination route are too numerous to encompass comprehensively in this report. Three examples will serve to illustrate the range of metallurgical possibil- ities considered. It will be obvious that certain schemes are less feasible technically and economically than others. In one separation scheme that probably belongs in the former category, recycled manganese is used as a cementation reagent to co-precipitate copper, nickel and cobalt (100). The precipitated metals are oxidation leached in ammonium carbonate to yield a solution from which copper and nickel are recovered by selective ion exchange and electrowinning. Cobalt recovery is by chemical precipitation from the raffinate. In another version, Kelex 100 liquid ion exchanger is employed to extract Cu, Ni and Co sequentially (123). Thus, copper is extracted from strongly acid solution and nickel and cobalt are extracted at the controlled pH ' s of 3.5 and 4.2, respectively. Other combinations of extractants have also been reported, including use of a trialkyl amine for selective cobalt extraction (29). A potential advantage of extraction by organic reagents of the liquid ion exchange variety is that stripping can be effected by sulfuric acid, leading to sulfate solutions of the metals. This is expecially advantageous in terms of cathode quality in the case of copper electrowinning. 2-11 2.3 SELECTION OF LIKELY PROCESSES Several excellent review articles have been published in the last few years describing the properties of nodules as well as some potential processing techniques, dictated by their composition, and the probable ranges of production levels and economics. Of particular interest, from a processing point of view, are summaries of the possibilities published by Agarwal et al. (3),Barbier (8), Cardwell (24), Faugeras (34), and Tinsley (123). All of these reviews point out that the most likely processing approaches involve smelting or hydrometallurgical processes including high or low temperature sulfuric acid leaching, ammonia /ammonium carbonate leaching following reduction, or hydrochloric acid leaching. The reviews by Agarwal et al. and Cardwell in particular, present fairly complete descriptions of the properties of nodules and economic considerations which dictate these process selections. Nevertheless, it is worthwhile presenting the rationale we have developed for arriving at the conclusion as to the most likely processing techniques. It is understood that the grade and ore reserves available will ultimately dictate the process and production level chosen, but, within limits, it is to be expected that a favorable ore can be made available for each process option. Then, the key decision from a processing point of view is whether or not to recover manganese as a primary process option or simply to build a so-called three-metals plant to recover copper, nickel, cobalt, and perhaps molybdenum. Putting aside marketing and economic considerations bearing on this decision for the moment, we will simply examine the processing impact of either decision. Excluding manganese, the value of nodules is as a nickeliferous ore with significant credits available from copper and cobalt. As such, it must compete with later itic ore which would typically contain about 1.5% nickel and 0.1% cobalt. With recoveries of 75% for nickel and 50% for cobalt, the returns from this ore will be of the order of $45-50/ton above mining costs, which are low. For nodules, however, mining and transportation costs are likely to be high, perhaps of the order of $20/ton (see section on Economics of Nodule Mining and Processing) so that, even at comparable capital and operating costs, returns will have to be of the order of $65-70/ton to compete with laterites. For the high valued 2.9% Ni/Cu/Co ore, recoveries of 70-75% of all metals would be required, with nearly complete recovery needed from the "average" ore. Even at $10/ton delivered nodules costs, recoveries of over 70% are required, and higher recoveries are obviously desirable to offset the higher risks inherent in the development of new technology. Energy costs and capital charges are key components in the costs of laterites processing, and it is not likely tliPt they will both be genfrally Lower for nodules processing. Thus, other costs, particularly process materials (make-up lixiviant) , must be carefully controlled to obtain favorable economics. The process must be essentially closed with respect to lixiviant use through effective regeneration/recycle or a special situation must exist with respect to the economics of supply and disposal of make-up and spent lixiviant. Since low temperature sulfuric acid leach processes (18) (37) (42) (44) show rather low metal recoveries even at long leach times, they are not likely candid- ates for economic processing routes. The same consideration applies to segregation roasting (49). High temperature sulfuric acid leaches, however, show good to excellent metals recoveries (20) (45) (83) (125) . While sulfuric acid consumption 2-12 may be as high as 0.3 lb/lb nodules, it is a readily available and relatively cheap lixiviant, and treatment and containment of the relatively small volumes of process wastes (excluding tailings) should not be too troublesome. Further- more, since the bases of this technology are practiced in laterites processing, at Moa Bay, an element of risk is removed which helps recommend high temperature sulfuric acid leaching as a likely processing alternative. Reduction leach processes based on ammonia/ ammonium carbonate as a lixiviant are also well known in laterites processing. Furthermore, they show good to excellent metals recoveries and are highly selective (9, 19, 43, 93, 132, 133). The newly developed hydrometallurgical , low temperature reduction/leach process (90, 120) is particularly interesting in that it combines operations and should save energy, which would help to offset the energy requirements involved in lixiviant recovery and recycle. In these terms, then, high temperature sulfuric acid and ammonia leach must be considered as likely processing alternatives for a three-metal plant. Basically the same considerations will apply in selecting probable process routes for processes which recover manganese as well as copper/nickel/cobalt. In this case, however, there is no analogous current practice that might guide tech- nology development. While potential revenues from many areas are high at present prices, a significant nodule production could cause sarious oversupply with concommitant price decreases. Both market and technology considerations suggest that these processes would be commercialized at a smaller scale than the three- metal plants. Smelting processes are candidates for routes involving manganese recovery because their inherently high energy requirements demand high revenues to offset these costs. Since technology exists for metals recovery from copper/nickel sulfide mattes, an element of risk is removed, and it is known that recovery of those metals can be high (13) (124) (116). If additional revenues from manganese production are indeed required, a simple reduction/segregation route is not likely to be viable (114) . While significant questions remain as to the purity of the manganese product and the overall process economics, a smelting approach cannot be rejected from consideration a priori. Both sulfating roast (20) (60) (79) (126) and sulfur dioxide reduction leach (20) (61) (64) processes are candidates for routes to manganese recovery since solubilization of all metals is high. However, problems arise in trying to construct integrated processes based on these approaches. As the metals are recovered from solution, the sulfate anion is converted to sulfuric acid. It is not likely that it, in turn, could be economically reduced to sulfur dioxide for recycle. It will be purged from the system as a relatively low strength, quite impure acid for sale for by-product credits. Disposal will likely require neutralization with large consumptions of base and the generation of large waste streams. The attraction of a "free" sulfating agent is illusory, since the sulfur requirements for a modest size nodules plant are so high as to require a site located adjacent to a very large power plant (5000mW) or copper smelter (100,000 TPY) which must be considered unlikely. Thus, sulfating systems are not likely to be viable process alternatives. The same considerations apply to schemes in which the lixiviant is ammonium sulfate (101) (102). 2-13 Potentially, the same objections could be raised with respect to systems employing chloride reduction roast/leach schemes, in spite of the acknowledged high recoveries of all the metals (25) (29) (41) (57) (77) (67) (125). In this case, however, the reduction reaction liberates a relatively pure, valuable by-product (chlorine) and the lixiviant (chloride ion) can be oxidized during metal recovery to produce additional chlorine or recovered as hydrogen chloride for recycle. It then may be possible, as a special circumstance, to arrange buy-back arrangements in HCI/CI2 with a manufacturer of chlorinated hydrocarbons adjacent to a chlor-alkali complex, although the tonnages involved would be quite large. The alternative is to buy large amounts of hydrocarbons of hydrogen directly to regenerate the hydrogen chloride. While this would involve known technology, it will not enhance process economics. Nevertheless, major objectives can be overcome, and chloride based processes must be considered where manganese produc- tion is a process objective. Finally, it should be noted that manganese recovery by physical means (bene- ficiation) or subsequent chemical processing of leach tails from the ammoniacal or high temperature sulfuric acid leaches would always remain a possibility. In summary, we conclude that there are five likely process options for the recovery of metal values from nodules : • For three-metal plants (copper/nickel/cobalt) either high temperature sulfuric acid leach or hydro or pyrometallurgical reduction leach processes using ammonia/ ammonium carbonate, both variations of known technology for laterites processing. « For four-metal plants (including manganese) either smelting processes using known technology for recovering metals from concentrated mattes or chloride reduction/leach processes, probably integrated with chlor- inated hydrocarbon production plants. 2-14 3.0 ECONOMICS OF NODULE MINING AND PROCESSING 3.1 INTRODUCTION In the past decade, a series of economic assessments has appeared in the rapidly growing literature on manganese nodule mining and processing. While there are wide variations among the estimates, factors of two being common, there is general agreement that manganese nodules offer potential returns sufficient to offset the significant risks associated with an ocean mining venture and the development of new technology. The objective of this study is to analyze the probable scale of operation of first generation land-based nodule processing plants. Since nodule mining and processing will have typical economies of scale, larger operations should reduce unit costs and increase margins. At the same time, however, even the ocean mining consortia face limits on the amounts of capital they can invest in such a venture. The most probable scale for first generation operations, therefore, is one at which unit costs have been lowered sufficiently to provide adequate margin, while capital requirements are within consortia financing capabilities. Such an analysis is, of necessity, subject to considerable uncertainty. In addition to the usual problem of evaluating future metal markets, the form, cost, and scale economics of nodule mining and processing technology are uncertain. A review of recent economic estimates concerning nodules, coupled with assumed analogies to land-based mining costs, yields a reasonably clear picture of the risks associated with various scales of operation. The final decision, however, will remain dependent on corporate philosophy. Parameters on which the economics of nodule operations will depend include: • extent of ore reserves • grade of ore • mining efficiency and rate • metallurgical recovery efficiency • capital and operating costs « metal markets and prices • legal and political conditions These factors are discussed in the following sections. 3.2 ORE RESERVES AND GRADES While manganese nodules appear to abound in significant areas of the world oceans, first generation commercial targets are likely to be located primarily in the near-equatorial North Pacific because of the apparent quality of the resource. Concentrations of nodules (population density) on the ocean floor vary widely, but average concentrations in major deposits are in the range of 2-3 lb/ft2 (1-2)*. ^Citations for Section 3 only are summarized in Section 3.9 and are not included in the process bibliography, Section 14. 3-1 The chemical composition of nodules varies widely, even within individual deposits, as is also the case with land-based ores. While a host of minerals is present in all nodules, commercial interest now centers on four: nickel, copper, cobalt, and manganese. Nickel is probably the key to ocean mining. Nickel concentrations of 1.1 to 1.6 percent appear to be prevalent in large areas of the North Pacific. Such grades are comparable to those of laterite nickel ores currently under considera- tion for new mining ventures (3). But nodules offer co-products that may signi- ficantly increase their metal values above those of new laterite ores. Copper is the principal co-product. Common concentrations in the North Pacific average 1.0-1.4 percent, somewhat superior to copper concentrations found in most domestic land -based ores now mined. Cobalt is essentially a by-product, found in relatively low concentrations of 0.15 to 0.25 percent when accompanied by high nickel and copper values. Nodule cobalt, however, can reportedly be recovered at about the same cost as nodule copper (4), making it a highly profit- able output at current prices. The economic attractiveness of manganese recovery is less clear. North Pacific nodules contain roughly 25 to 35 percent Mn which, if recovered, could double revenues from an ocean mining operation at current prices. Yet only the Deepsea Ventures consortium has announced plans to recover high-purity manganese metal from first generation operations, while Ocean Management apparently will recover lower-purity manganese products (5). At least some interest has been shown in recovering additional minerals from nodules, most notably molybdenum and zinc (6). Their recovery, however, will depend upon detailed assays of a specific ocean mining site and should affect overall economics only marginally. To summarize, ore reserves represented by manganese nodules appear to be huge and occur in sufficiently concentrated deposits to support orderly mining from well-defined claim or lease areas. Combined nickel-copper-cobalt grades of 2.5 percent offer superior revenues to most available laterite nickel deposits, with manganese a potential source of additional income. 3.3 MINING AND PROCESSING TECHNOLOGY The major ocean mining consortia have expended some $25 to $30 million each on technology development for nodules operations (7) . Research and development efforts have focused on testing components of mining systems and small-scale metallurgical processing plants. The leading groups appear ready to enter the final pre-production stage of development in which the remaining technological uncertainties would be resolved. Development of an adequate mining system represents a formidable task, since selectively lifting the manganese nodules some three miles from the deep ocean floor to its surface represents entirely new technology. Two main types of mining systems are being developed by industry to mine deep ocean manganese nodules: hydraulic and mechanical. At present, hydraulic systems seem to be favored by U.S. industry over the mechanical system. Both types of systems have been tested on a pilot scale (8) , with additional pilot scale as well as prototype testing scheduled during the next few years. 3-2 Hydraulic mining systems are designed to recovery nodules by creating an upward flow of seawater and nodules from a collector on the seafloor, through a pipe to a surface mining ship. The mechanical system under develop- ment, the CLB system, consists of a series of buckets attached to a continuous line that will travel from one mining ship to the seafloor, along the bottom up to a second mining ship and over to the first ship. Dredges and pipelines must be linked to surface vessels capable of extremely accurate navigation and maneuvering. Some conceptual designs include a deep submergence vessel for controlling the operation or for maintenance and repair (9) , the temporary storage of mined ore on barges that would double in transporting nodules to port behind ocean-going tugs (10) , or simply expanding the control ships to include bulk storage capacity which would take advantage of economies of scale in ship-building. Alternative metallurgical processes for treating the mined nodules are thoroughly discussed elsewhere in this report. While processing may be pyrometal- lurgical or hydrometallurgical , there is no reason to believe, at this point, that the economics of one or another route are inherently favored. A key discrim- inator for capacity and economics, however, will be the decision on recovery of manganese. 3.4 CAPITAL AND OPERATING COSTS Although many estimates of the capital and operating costs of nodule mining and processing operations have been prepared, they follow no consistent set of assumptions or forecasts. Tables 3.1 and 3.2 summarize estimates for hydraulic dredge mining systems and hydrometallurgical processing systems ranging in size from one to three million dry tons per year. Figures 3.1 and 3.2 display the range of total cost estimates for these scales of operation. All estimates have been deflated to mid-1976 dollars. Despite major variations among estimates, something of a pattern emerges in Figures 3.1 and 3.2. At a scale of one million dry tons per year, capital cost estimates range roughly from $175 to $250 million dollars. At three million tpy, estimates vary from $400 to nearly $800 million. With the exception of the last figure, these ranges are consistent with a capital cost-capacity exponent of about 0.8. Operating cost estimates vary even more widely, but display the same trend toward economies of scale. At one million tpy the average operating cost estimage is about $70 per dry ton while at 3 million tons/year, the average decreases to about $45 per dry ton. Table 3.1 Nodule Operation* Capi tal Req uirements : Estimates from the Literature (mi llion 1976 $) Rothstein & Sorenson Moncrieff & Author** Kaufman Tinsley CMS & Mead ?male-Adams Wright Date 1973 1976 1975 1968 1974 1976 Scale (million dt py) 1.0 1.1 1.5 1.8 3.0 3.0 R&D - 56-62 40 - 70-170 75-150 Mining - 44 29 251 - - Transport - 4 31 25 - - Processing - 80-115 88 84 - - Other - 15-17 - - - - Total 171-316 200-240 188 360 415-780 500-750 Assumes dredge with hydraulic lift, hydrometallurgical process, 3-metal recovery, **Sources: Ref. 7, 10, 12 through 16. 3-3 CAPITAL INVESTMENT Million 1976$ TOTAL OPERATING COSTS 1976 $/DST 600 • 500 400 300 200 100 L / / / / / / / 120 100 t 12 3 MILLION DRY STPY FIGURE 3.1: CAPITAL COST ESTIMATES 80 r I ! 60 \ j 40 i- 20 \ 12 3 MILLION DRY STPY FIG. 3. 2: OPERATING COST ESTIMATES 1 . Kaufman 2. Tinsley 3. U.S. Commission on Marine Sciences, Eng . and Resources 4. Sorensen and Mead 5. Moncrieff & Smale-Adams 6. Wright 3-4 Table 3.2 Nodule Operat ing Costs .* Estimates from the Literature (1976 $ per dry ton) Sorensen Moncrief f & Author** Kaufman Tinsley CMS & Mead Smale-Adams t^right Date 1970 1976 1975 1968 1974 1976 Scale (million dtpy) 1.0 1.1 1.5 1.8 3.0 3.0 Mining - 9 6 28 - - Transport - 7 6 17 - - Process - 19-25 32 42 - - Other - 16-20 - - - - Total 32-62 50-62 44 87 47-77 40-55 To provide a clearer picture of capacity-cost relationships, Tinsley 's latest estimates and supporting data (14) were combined with judgments concerning likely economies of scale in each component of mining and processing. The resulting estimates are shown in Tables 3.3 and 3.4 3.5 MINING The mining cost assessment (table 3.3) assumes that Glomar Explorer - type ships will be employed, with expanded bulk storage capacity for seven days produc- tion. The one million dry metric tpy ship, taken from Tinsley, is based on a 50,000 dwt vessel with 33,000 dwt for storage (7 days at 4,700 tons per day). For the larger scale operations, costs were estimated for both one- and two-ship operations. The one-ship estimates provide an indication of cost sensitivity only; it is doubtful whether a single mining ship and dredge with more than a two million tpy capacity could be built. Hull cost estimates are based on Tinsley 's reported foreign construction costs for new bulk carriers (extrapolated for the 85,000 and 120,000 dwt vessels), while costs of modifications, navigation equipment, etc., are assumed constant for all sizes of mining ships. Dredging equipment and bottom miner costs are scaled upward from Tinsley 's one million tpy estimate using an eight-tenths capacity exponent. Ore transportation costs are based on an assumed 4,000 mile round trip from a West Coast port to the Pacific mining site. Again, Tinsley 's data on ocean bulk transport rates were used, with the assumption that all carriers will be leased on a long-term basis to minimize investment requirements. Port facility costs are assumed conservatively to be proportional to ship capacity. Exploration and R&D costs are assumed constant for all scales of operation and working capital requirements are estimated at 8 percent of fixed capital. The resulting total capital investment estimates for the mining system range from about $100 million for a one million dry metric ton per year operation to $172 million for three million dtpy (Ajith two mining ships. Scaling up the mining ship and dredges to two or three million dry metric tons per year could significantly reduce capital costs per ton if technologically feasible. Assumes dredge with hydraulic lift, hydrometallurgical process, 3-metal recovery . **Sources: Ref. 7, 10, 12 through 16. 3-5 Table 3.3 Capital Investment (million 1976 $) Fixed : Ships^ Airlift and Pipe'^ Bottom Miner Port Facilities Total Fixed Exploration and R&D Working Capital^ Total Investment Nodule M ining Cost Estimates dt Scale of Operati ons (million py) b} 2 b2 ^ 3 b2 36.0 72.0 45.0 84.0 51.0 6.0 12.0 10.0 16.5 14.5 2.0 4.0 3.5 6.0 5.0 4.0 4.0 8.0 6.0 12.0 48.0 92.0 66.5 112.5 82.5 50.0 50.0 50.0 50.0 50.0 4.0 7.0 5.5 9.0 6.5 102.0 149.0 122.0 171.5 139.0 Operating Costs ($ per wet ton) Mining :6 Power 4.65 4.65 4.65 4 65 4.65 Bunker 0.35 0.35 0.35 35 0.35 Wages 1.53 1.53 1.08 1 25 0.88 Insurance^ 0.77 0.74 0.54 61 0.44 Supplies 0.11 0.11 0.11 11 0.11 Overhead & Misc : . 0.27 0.27 0.27 27 0.27 Capital Charges ,8 7.38 7.08 5.12 5 77 4.23 Total Mining $15.06 $14.73 $12.12 $13 01 $10.93 Transportation & Handling 6.80 6.80 4.60 5 10 4.10 Total Cost/Wet Ton $21.86 $21.53 $16.72 $18 11 $15.03 Total Cost/Dry Ton $28.40 $28.00 $21.70 $23 50 $19.50 Not es to Table 3.3 for 2 million tpy for 3 million tpy 1. Two ship operation. 2. One ship operation. 3. Mining ship sizes: 1 50,000 dwt for 1 million tpy 50,000 dwt or 85,000 dwt 65,000 dwt or 120,000 dwt 4. Airlift, pipe and bottom miner scaled with 0.8 capacity exponent. 5. Working capital = 8% of fixed capital. 6. Mining of 1.3 to 3.9 million wet metric tpy, 1.0 to 3.0 million dry. Power at 1.5 hp per ton per day for hoisting, 3.0 hp per ton per day for propulsion, 0.5 hp per ton per day (net) for pipe handling, 0.2 hp per ton per day for ore transfer. Wages scaled with 0.5 capacity exponent. 7. Insurance at 2.1% of fixed capital per year. 8. Capital charges at 20% of fixed capital, covering depreciation, maintenance, taxes, interest, etc. 3-6 Operating cost estimates for mining are presented in Table 3.3. On a per-ton basis, wages, insura.ice, capital charges and transport costs are assumed to depend on scale. Mini'ig wages are estimated with a 0.5 capacity exponent, while insurance and capital cnarges are constant fractions of fixed capital investment. Ore transport costs are based on Tinsley's data for ocean freight rates, assuming freighter sizes increase with mining capacity per ship. Cost estimates for matallurgical processing are summarized in Table 3.4. The basis for these calculations is Tinsley's estimate of $115 million for a three metal process plant using an ammonia leach. This process is essentially an adaptation of known technology (laterites processing using the Caron process) for which economics are well known. Thus, the estimate should be within + 30% of the likely cost of the plant and should also approximately represent the costs of three metal recovery plants using different process schemes. Estimates of the capital costs for processing at higher rates were made by using an 0.8 capacity exponent. R&D costs are assumed constant across scales while working capital is set at 8 percent of plant costs. Table 3.4 Three-Metal Nodule Processing Cost Estimates Scale of Operation (million itpy) 111 115 200 277 Capital Investment Fixed: Plant R&D Working Capital-*- Total Investment Operating Costs (per dry ton) Supplies Utilities Labor2 Capital Charges-^ 50 9 50 16 50 22 174 266 349 4.9 4.9 4.9 13.0 13.0 13.0 1.90 1.25 0.98 23.0 20.0 18.5 42.80 39.15 35.38 Working capital = 8% fixed capital ^Wages scaled with 0.4 capacity exponent ^Capital charges = 20% fixed capital Process plant operating cost estimates are also adapted from Tinsley, but with capital charges included to cover maintenance, depreciation taxes, etc. Supplies and fuel costs are assumed to be insensitive to size, while labor costs are scaled using a 0.4 capacity exponent. The advantages of large scale nodule mining and processing operations are apparent. At one million dtpy, capital requirements amount to some $275/ton, while combined ore and process operating costs exceed $70/ton. At three million dtpy, however, the necessary investment is less than $175/ton of capacity, while operating costs could drop to around $60/ton. While larger scale operations should yield further cost reductions, it may be difficult to satisfy capital requirements much greater than half a billion dollars, even by a consortium, on a first generation operation. 3-7 The costs of nodules processing in a four metal plant, including manganese, are more difficult to define since there is no current production scheme from which to draw analogies. With additional revenues, however, and considering Mn market impact problems, it is likely that first generation, four-metal plants will be considerably smaller than three-metal plants. It should be noted that, in principle, the option always exists to add on a manganese recovery from tailings step to an existing three -metals plant. In that case, any production scale desirable could be run, and the venture's success would depend on the processor's ability to upgrade a "free" ore in competition with currently exploited resources. However, other than a proposal to recover manganese by reducing slags from smelting operations, no information on technology appropriate to such a purpose has been published, so this option will not be considered further here. The impact of manganese recovery on nodules plant processing costs is, as a first approximation, to increase only the metals recovery and utilities component of the capital costs by increasing total metals recovery about ten times. But since these components represent only about 20% of the total cost of a three- metal plant, and since economies of scale are possible, the total capital required for a four metal plant should be only double that required for a three-metal plant of equal capacity. By the same logic, operating costs per ton of nodules will also approximately double. While materials and supplies costs should increase moderately, capital charges, representing half the costs double, and utilities and labor costs will increase significantly by virture of the greatly increased metals production. For orientation purposes, then, we will assume a factor of two is appropriate for reflecting the increased costs of manganese recovery. The economies of scale for this case are summarised in Tables 3.5 and 3.6- 3.6 METAL MARKETS AND PRICES The impact of nodule operations on metal markets and the ability of the markets to support nodule mining have been subjects of considerable disagree- ment. Many forecasts of future demand and supply of the major nodule metals are available and need not be supplemented here. More important are the apparent trends in metal prices and their influence on investment decisions. 3.6.1 Nickel Long term nickel demand has been forecast to grow in excess of 6 percent per year (17). Assuming such a growth rate occurs, world production will have to increase some 675 thousand tons per year by 1985 to balance the market. This gap would appear to offer room for both nodule operations and new land-based mines. In the face of expected tight supplies, nickel prices are expected to rise steadily in real terms — perhaps to a level of $2.50-$2.60 by 1985 (in 1976 dollars) even if nodules achieve market penetrations of several percent (7). Such price levels, in fact, would probably be required to develop more later ites capacity in unstable and inaccessible areas. 3-8 Table 3.5 Four-Metal Nodule Mining Cost Estimates (million 1976 $) Capital Investment Fixed : Ship Airlift and Pipe Bottom Miner Port Facilities Total Fixed Exploration and R&D Working Capital (8%) Scale of Op eration (Mi llion dtpy) 0.5 1.0 30 4 2 2 36 6 2 4 38 48 50 50 3 4 Total Investment 91 102 Operating Costs (per wet ton) Mining ; Power Bunker Wages Insurance Supplies Overhead & Misc. Capital Charges Total Mining Transport Handling Total Cost/Wet Ton Total Cost/Dry Ton 4.65 4.65 0.35 0.35 2.16 1.53 1.23 0.77 0.11 0.11 0.27 0.27 11.69 7.38 20 .46 7 .00 1 .10 15 06 5 70 1 10 28.56 21.86 37.1 28.4 Table 3.6 Four-Metal Nodule Processing Cost Estimates* (million 1976 $) Scale of Operation (million dtpy) Capital Investment . 0. 5 1 .0 Fixed: Plant R&D Working Capital Total Investment Operating Costs ($ per dry ton) Supplies Utilities Labor Capital Charges Total 135 235 50 50 11 196 19 304 5.7 5.7 26.0 26.0 5.7 4.3 54.0 47.0 91.4 83.0 *See notes to Table 3.3 3-9 3.6.2 Copper World consumption of refined copper is now around 10 million tons per year and is growing at some 4-5 percent per year (18). Most analysts feel that copper is in over-supply and will continue to be so for several years. New capacity that will be required in the 1980 's or sooner, however, cannot be supplied at current prices. Forecasts of $0.80 to $1.00 per pound of refined copper in the long run are not unusual (7). In any case, the copper market will be left unperturbed by nodule operations in the foreseeable future. A 10 million ton per year nodule industry could supply only about 1.3 percent of current copper demand. The principal question, then, is how much an independent rise in copper prices could improve nodule revenues . 3.6.3 Cobalt The market for cobalt almost certainly would be affected by a growing nodule industry. A single three million tons per year nodule operation could supply nearly 25 percent of 1974 world production. By 1985, nodules could supply from 10 to 50 percent of world consumption, even with relatively high demand increases in the interim, and the most widely held expectation is that there will be a proportionate drop in price. The magnitude of such a decrease, however, would depend on the extent to which cobalt could be substituted for nickel as supplies are increased. While it is possible that cobalt prices will fall to or below that of nickel, Wright assumes a 1985 price of 25 percent above current levels in her assessment of nodule economics (7) . 3.6.4 Manganese The market for pure manganese metal is insufficient to support ocean mining production. A single one million tpy operation could produce several times the current world consumption of manganese metal. Nodule manganese, however, should be competitive in the 5-7 million tpy free world ferromanganese market. Deepsea Ventures reportedly will attempt to sell in the higher-priced metal and medium-to-low carbon ferromanganese markets (5) , while OMI will produce either a synthetic ore or ferromanganese exclusively. Sustaining a premium price above that of standard high-carbon ferromanganese (now about 25 cents per pound) may be difficult. Demand for ferromanganese depends largely on steel production and should sustain moderate growth in the coming decades. The availability of nodule manganese, however, should eliminate significant price increases outside normal business-cycle fluctuations, and could encourage price reductions. Most investigators conclude from reviews of these data that nickel and copper prices are subject to upward pressures that should continue largely unabated despite the availability of a nodule production. While cobalt prices may fall, they will not significantly alter nodule economics. The price and market sit- uations for ferromanganese should be stable, unless an increase in interest in ocean mining manganese production causes moderate price decreases. 3-10 3.7 POTENTIAL REVENUES AND PROFITS The potential profitability of nodule operations will be determined by the capital and operating costs of the venture at a given scale of operation, the magnitude of which has already been estimated. Revenues, then, depend on metals prices previously considered, and ore grades. The metal contents of nodule ores are highly variable, and while average grades over a 20 year mine life might not exceed 2.5 percent for copper/nickel/cobalt, combined grades could reach 3.0 percent in first-generation operations focused on the best available mining sites. There is likely to be some fluctuation in grade of ore over project life, but the uncertanties involved reduce the utility of discounted cash flow or similar dynamic financial analyses for this brief review. Table 3.7 indicates assumed high range and low range grades of nodule ores, coupled with forecast prices and process efficiencies to determine net revenues per metric ton of ore for a three-metal plant. Table 3.8 summarizes the overall financial status of three-metal nodule operations ranging from one to three million dty in scale. Both high range and low range revenues are employed for comparison. Table 3.7 Nodule Operation Rev( snues High Range Low Range Price* Re covery Net Value (per ton) Mineral Grade (%) Grade (%) (per lb) Efficiency High Range Low Range Ni 1.5 1.3 $2.50 90% $ 74 $ 64 Cu 1.3 1.1 $0.85 90% $ 22 $ 19 Co 0.25 0.2 $3.00 60% $ 10 $ 8 Three-Metal Total $106 $ 91 Mn 25.0 25.0 $0.25 90% $124 $124 Four-Metal Total $230 $215 Table 3.8 Potential Profits of a Three-Metal Nodule Processing Operation (million 1976 $) High Range Total Investment Revenue Ore Cost Process Cost Margin After Tax Profit ROI Low Range Revenue Margin After Tax Profit ROI Scale of Operation ('""■•"llion dty) 12 3 ,A* ,*>VA a'" ** 276 415 388 521 488 106 212 212 318 318 28 56 44 71 59 43 78 78 106 106 35 78 90 141 153 18 39 45 72 77 6.5% 9.4% 11.6% 13.8% 15.8% 91 182 182 273 273 20 48 60 96 108 10 24 30 48 54 3.6% 5.8% 7.7% 9.2% 11.1% ^Assumed future prices based on common forecasts. See text for discussion, Two-ship mining operation '"'"'' One-ship mining operation 3-11 Simple return-on-investment calculations reveal that a three-metal operation becomes attractive only at a 2-3 million tpy scale. High range simple returns of 10-15 percent, when supplemented by significant depreciation, imply an ade- quate discounted cash flow for such investments. The low range ROIs suggest that 2 million tpy may be the cutoff point below which investments will not be made. These calculations also assume that no depletion allowance will be made. The same estimates for four -metal operations are shown in Table 3.9. The simple returns are quite adequate at a one million tpy level and marginal at 0.5 million tpy. Clearly, however, the success of a four-metai operation will depend heavily on stable manganese prices and successful market penetration. Table 3.9 Potential Profits of a Four -Metal Nodule Processing Operation (million 1976 $) Scale of Operation (million dtpy) 0.5 1.0 287 406 115 230 19 28 45 83 51 119 26 60 9.1% 14.8% 108 215 44 104 22 52 7.7% 12.8% High Range , Total Investment Revenue Ore Cost Process Cost Margin After Tax Profit ROT Low Range Revenue Margin After Tax Profit ROI These levels of profitability might be viewed as barely adequate to justify the assumption of the risks involved. The trend in new land-based mines (particularly nickel), however, is toward extreme infrastructure requirements which greatly inflate capital investment. Recent reports of land-based nickel investments suggest a capital cost in the range of $6-$8 per pound of nickel equivalent (7). Nodule operations, at roughly $5 per pound nickel equivalent (in a three-metal, 3 million tpy operation), compare quite favorably. Operating costs "per pound appear to be similar for new land-based and ocean mining oper- ations. Prices may therefore be anticipated to rise until the lowest-cost new production units achieve acceptable returns. Further, the intangible benefits of a domestic processing plant versus heavy investment located in politically unstable developing countries should not be ignored. This assumes, of course, that the international Law of the Sea and negotiations or the enactment of U.S. domestic legislation result in the establishment of a stable legal regime which reduces political risks to the investment. Finally, second generation nodule operations are likely to be considerably more profitable. Research and development requirements should be greatly reduced, learning-curve effects will reduce operating costs, and scales of operation could exceed 5 million tpy. While these possibilities have no direct bearing on the initial investment decisions, they no doubt are one source of current interest in nodule mining. 3-12 3.8 CONCLUSIONS The first three-metal nodule operations are likely to be sized at about 3 million dry tons per year. This scale offers adequate potential returns and could break even, despite major unplanned cost increases. Smaller-sized operations have marginal profit potentials, at best. We have assumed in our research and development cost estimates that major pilot scale efforts to prove mining and processing technology will be undertaken prior to the first commer- cial-scale investments. On this basis, initial commercial operations less than 2 million tons per year make little economic sense. Four-metal nodule operations could be much smaller, in the range of about 0.5 to 1.0 million tons per year. Because of manganese revenues, these sizes offer potential profits in an acceptable range. Equally important, capital requirements per ton of capacity are significantly higher than those of a three-metal operation. Marketing uncertainties also argue for the smallest possible investment. It should be stressed, however, that cost estimates in this study for the four-metal operation are no more than rough approximations. 3.9 ECONOMIC CITATIONS 1 Mero , J.L., The Mineral Resources of the Sea , Elsevier Publishing Co., New York, 1965. 2 Mero, J.L., "Potential Economic Value of Ocean Floor Manganese Nodule Deposits", in D.R. Horn, ed . Ferromanganese Deposits on the Ocean Floor , Conference, Arden House, Harriman, N.Y., Lamont-Doherty Geo- logical Observatory, Columbia Univ. and IDOE, NSF, 197 2. 3 Boin, U. and E. Muller, "Economic Aspects of Manganese Nodule Deep Ocean Mining", Metallgesellschaf t , A.G. Review of Activities, Edition 18, 1975. 4 Tinsley, C.R., "Economics of Deep Ocean Resources - A Question of Manganese or No Manganese", Mining Engineering , April 197 5. 5 Tinsley, C.R., "A New Picture Emerges in Deep Ocean Mining", Mining Engineering , April 1976. 6 Cardwell , P.H., "Extractive Metallurgy of Ocean Nodules", Mining Congress Journal , November 1973. 7 Wright, R.L., "Ocean Mining: An Economic Evaluation", Professional Staff Study, Ocean Mining Administration, Washington, May 1976, p. 3. 8 Anon., "Ocean Mining Comes of Age", Oceanology International , December 1971. 9 Swan, D.B., "The Potential of Manganese Nodules as a Future Mineral Resource", Marine Technology , January 1974. 10 Sorensen, P.E. and W.J. Mean, "A Cost-Benefit Analysis of Ocean Mineral Resource Development: The Case of Manganese Nodules", J. Agricultural Economics , December 1968. 11 Agarwal, J.C, N. Beecher, D.S. Davies, G.L. Hubred, V.K. Kakaria and R.N. Kust , "Processing of Ocean Nodules: A Technical and Economic Appraisal", Journal of Metals , April 1976. 3-13 12 Kaufman, A.J. "A Survey of the Economics of Ocean Mining", Mar ine Technological Society Journal , Vol. 4, No. 4, 1970. 13 Kaufman, R. and A.J. Rothstein, "The Approaching Maturity of Deep Ocean Mining - The Pace Quickens", Mining Engineering , April 1974. 14 Tinsley, C.R. , "The Future Markets for Nodule Metals", Proceedings, A. I.M.E. /M.M. I. J. Conference, Denver, September 1976. 15 U.S. Commission on Marine Sciences, Engineering, and Resources, Marine Resources and Legal Political Arrangements for their Development , Vol. 3, U.S. Government Printing Office, Washington, D.C., 1969. 16 Moncrieff, A.G. and K.B. Smale-Adams, "The Economics of First Generation Manganese Nodule Operations", Mining Congress Journal , December 1974. 17 Sahl, N.G., "Nickel ~ 1975 Slump May Be Followed By Long Term Trend Toward Short Supply", Engineering and Mining Journal , March 1976 18 American Bureau of Metal Statistics, Non-Ferrous Metal Data , 1975, Apex Press, N.Y. , 1976 3-14 4.0 GENERAL STUDY APPROACH AND ASSUMPTIONS 4 . 1 APPROACH The fundamental step in describing manganese nodule treatment plants based on the processes identified in Chapter 2 was the development of process flow sheets and process material and energy balances. Once these data were developed for each process, description and characterization of plant inputs and outputs, the physical and chemical characteristics of plant wastes, the distribution of toxic materials, the type, magnitude and frequency of accidental releases of hazardous materials, plant water, power, land and manpower require- ments and other factors needed to assess the probable impacts from each type of plant could be developed. Assumed production rates were established in accord- ance with the conclusions of the economic analysis presented in Chapter 3. For four -metal plants (i.e., those recovering manganese), the production rate was indicated to be about 30% of that for a three-metal plant (i.e. , recovering nickel, copper and cobalt but not manganese). A minimum processing rate was chosen to provide the necessary revenues for the target return on investment. Published technical literature and patents provided much of the data needed to develop the basic flow-sheets and material and energy balances. The literature was reviewed in some detail to glean any information which might be available on each process. Of major interest was information bearing on rate processes (reduction, extraction, liquid/solid separation, heat exchange fouling, etc.); equilibrium relationships (distribution of metals and impurities in slag and matte phases, vapor/liquid equilibrium in tailings stripping, etc.) and their relationship to processing conditions such as temperature, pressure, pH; and required residence times, yields, etc. from pertinent unit operations. The amount of pertinent data available in the open literature varied for each process. Where data needed to develop the process configuration and material and energy balance were not available, the needed criteria were generated on the basis of engineering judgement and assumptions, knowledge of . the state-of-the-art and/or a knowledge of analogous processes. The following paragraphs introduce general assumptions and judgements made during the course of this study. The following five sections of this report summarize the important characteristics of nodule processing plants designed around the five process schemes identified in Chapter 2. Process-specific assumptions made during the study are discussed in the appropriate sections . 4.2 ASSUMPTIONS 4.2.1 Manganese Nodules Composition It has been assumed that the nodules will be delivered to the plant via slurry pipeline from the port facilities and will be stored wet. Slurry water will be decanted and returned to the port. The assumed nodules composition for purposes of this study was a composite of published information sunmiarized in Table 4.1. In most instances, the condition of the nodules before drying and the extent of moisture removal during drying were not clearly specified, so that the assumed water content of 21% had an uncertain basis. The most likely basis, however, is for nodules which had been air dried, so that 40% is a more realistic water content for reclaimed wet nodules. In particular, the study assumption is that the nodules, as reclaimed from wet storage, will include 40% total water, distributed as follows: surface (i.e., pellicular and interstitial) water, 5%: pore water, 25%; chemically bound water, 10%. The 25% pore water corresponds to 48% nodules porosity if the specific gravity of the matrix is taken to be 2.8 4-1 ro m CO o o u"! o O .H vO O O O in 00 iH Lo 0^ 00 1— I O r-^ rsi O tH O O iri I o o o O o O -H O CNJ CM 00 O O CNi o a^ S-i C/3 i-1 00 m 00 CM o •< vD csi in vO vO tH ~3- r^ o ro ^ o o> 00 • CN • iH • O •■ ro a> -H o O -H rH in 00 ro ro rH O O Csl vO O O cN 00 in vO r-~ O O O (Nl ,H ro VD O cr\ H r~~ ~* r-^ rH O O O " •- o CO in 00 in iH c^ in O 00 o o c a^ iH in r~~ CO CO o cNi in csi ,H o- - i iH -H iH in Csl CO en Tl CO in in • • • • r-i C/1 o O CM e d CN sD o en d S-i CD i-H P. fi in CN (U cn 00 rH i-H B d B d e d •H d c o C •H e d d d d cu fi CJ •H e •H cu •H Id C 0) c Fi CO c o 4= u •H cn d •H •H •H •H C 4-) iH i-l d •H 4-1 d C T5 TD U d e c o d QJ ■H u a d >-l U) •H Tl c X) fi Cfl iH CU OJ •H C c •H 43 Cfl QJ •H e cd o J3 •1-1 c R •H (fl M-l crt c crt cfl Oi) C Cfl -:^ O- a iH 0) V-l >srH > ti •H 4-1 2 l-l M -o M) d 1—1 n iH .H 4-1 tH cn 4J C U C J3 CJ Cu C rH cn i-l 4-1 Vj rH rH rH TJ C 4-1 c o CD o cfl .— 1 •H 4:: d -C crt •H cfl x; Cfl Vj •H •H Cfl >-i 4-) 4-) •H Cfl •H tfl •H c e PQ U CAl :^ <: C/3 CL, CO U P- C/3 H > C_) s 1— 1 U 2 u N CJ5 < C/D >-' M S Pl4 C/J u H <: P3 CJ oc ' < 05 Ph 1-i cfl a -H to U ^ U c/5 H > C QJ 2 f^ •H d C cfl cn (-1 Z cj> M CJ -< O T3 bO T3 C J3 s pi4 <; CI c/3 to 4-: I H •H cn o o u to QJ i-H O 2; N > u CO O 0) pa 3 to 0) O O 00 m u cu M iH 13 ^ M • y, O 03 13 •H 3 H CO O M O CM O est CO in O o CN O CNI v^ CM rH r-- r-- r-- 1 1 t 1 1 1 o o o rH iH iH ro C» ^ CTi o iH rH . 3 ^ 0) e Cfl 3 Vj -h u B 3 4:: •H 3 .H 3 4J n) •H j-J 13 13 rH 13 g C 3 M c •H r-\ )-i ct3 cfl U] QJ cd ctt S-J QJ ^ (1) •H a pa hJ M S H kJ pq (U 1 M C nj cti l-i 3 M)i-IJ2.HO CN cqHJM<;5CHP-ipQ»_3K CO (U VI 3 4-J CO cn )-j •H dJ Pu 6^ e 0) vO 4-1 -c S ■H .-H ^ > >. rH MH (U > •H »■ 4J S-l cti CU rH +J OJ CO U > J-i B-« ct) o 3 C -H •H t^ CTJ S-i 4-i 13 3 !>^ V XI CD •U rH J3 13 CO (U > B 3 6 cn oj cn 5-1 CO 13 3 3 q; co cu Xi cn OJ OJ J-i > CO Cu ,3 0) cn rH cu ^ rH -H 3 cn XI cn 0) 3 cj CJ 13 CO Q) B 3 ■H -H CO rC ^ ^ a -H cu 5: Vj 13 « 0) ■u 3 cu -H tS CO W 3 • u CU ■U S-l CO CU s J-1 to 0) & a co cn H-i 3 C M rH 3 ex •H cn 4-) Vj •H U-l OJ c -u 00 CO •H 0) S > a •H CU cn 3 CO « rH 14H cn CJ U X 3 hJ W CO Ik + 4-3 or to 50% porosity if the mean matrix specific gravity is taken to be 3.0. This is a reasonable result in terms of published porosities. The content of all elements in the wet nodules, then, will be lower than for the air-dried basis presented in Table 4.1. The nodules composition, based on 40% water content, is summarized in Tables 4.2-4.4, where the elements are also classi- fied according to their value, toxicity and abundance. It should be emphasized that there is no typical or representative manganese nodule composition. The best that one can hope for is that the study-basis compo- sition is reasonable and self-consistent. That the assumed element precentages broadly meet these criteria is shown by the following considerations. The assumed nodules porosity, filled with standard sea water (1.9% by weight of chloride), would hold chloride in an amount equal to 0.5% of the total nodules weight, as compared with 0.64% shown in Table 4.3. Again, referring to Table 4.5, if the source of sulfur in the nodules were exclusively sea water (sulfate is the second most abundant anion in sea water) , the percent of sulfur in the nodules for the assumed 0.64% chloride would be 0.03%. The difference between this figure and the assumed 0.08% sulfur could be reasonably reconciled by assuming the remainder of the sulfur to be present as the very sparingly soluble barite (BaSO^, barium sulfate). This would account for a nodules barium content of 0.13%, the additional 0.2% Ba (Table 4.4) most probably being associated with the clay minerals. Additionally, if all of the identified constituents, except for chloride and equivalent sodium, are taken formally to be present as their highest stable oxides, the total solids content of the nodules (without surface water) is 69.7%, in reasonable agreement with the assumed 60/0.95 =63.2%. Of the calculated 69.7% solids content, the major and value metals and the innocuous nonminor elements, all with combined oxygen, account for 68.5%. The remaining 1.2% is made up formally of the oxides of some twenty minor and trace constituents. The identified members of the major categories of elements in Pacific Ocean nodules are displayed in the summary in Table 4.6. The innocuous minor elements are either inherently nontoxic or are not considered toxic at the low levels in which they occur in the nodules. Further, as a group, they constitute less than 0.3% of the weight of the nodules and so do not affect mass flows within the precision of the estimation of those quantities. Accordingly, the innocuous minor elements are not considered further. '^•2.2 Selection of Potentially Toxic or Hazardous Constituents Undisturbed manganese nodules are manifestly benign. Their ingredients are all derivable from substances dissolved in or suspended in the oceans. The presence of potentially toxic and hazardous elements in the nodules can only be of concern if they become concentrated in certain altered chemical forms. Some examples are release of Ni3S2 aerosols, volatile carbonyls, greater than trace concentrations of cadmium in effluent solutions, etc. These occurrences would presumably be guarded against in the processing of manganese nodules as is done in normal metallurgical practice. 4-4 S-i 0) w 4-1 w ^2 hJ t3 (1) o CJ o 2 i-i-i 1-1 ^_^ ^D Z 3 c o> <; en Kl O w v^^ o U-l o O CX) u 0) CX) 1— 1 > 3 fn •H U O M cn •H g CO 4-1 < a 4-1 c o CJ o o Pi CN o ,, --^ <)• •H cn n) 4J o o o s ■-' o - - I ■u u •H & < CO Pm 3 >> n • z o 1 o z o OJ •H cn a) Pi 3 P .H O ,, O X u CO w o •H CO CO 4-i M M i ■U C o 1— 1 ■ ■ CO VO i 4J •H O-l ! O H H O (^^l 00 to .H CO CNj s -H rH vO 00 i CO t^ X sc 2 iH o ■H ro ^ U CN r^ (y\ LTl in CM C3N ro en • • • LO Ni rsi CT> m CT\ CO \0 O 2 CN CO C o •• •• ■U J-> 4-J c c •- nj CO & iH rH O Oh O- 1^ --( rH CO CO iH -U 4-> CO (U 0) o H ■«■ en 4-6 cn o 42 •H CN s cn M I— 1 - CN s m o CO 00 : hj CN i-H O CO o 1 CM CJ^ CTn O c-si O O - Fluorine 1.4 Nitrogen 0.01-0.7 Aluminum 0.5 Rubidium 0.2 Lithium 0.1 Phosphorus 0.001-0.10 ->■ Barium 0.05 Iodine 0.05 ->■ Arsenic 0.01-0.02 Iron 0.002-0.02 -»- Manganese 0.001-0.01 Element -> Copper ->- Zinc -^ Lead Selenium Cesium Uranium Molybdenum Thorium Cerium Silver -> Vanadium -> Lanthanum Yttrium -> Nickel Scandium -> Mercury Gold Radium ->- Cadmium -^ Chromium ^ Cobalt Tin mg/kg (ppm ) 0.001-0.01 0.005 0.004 0.004 0.002 0.0015 0.0005 < 0.0005 0.0004 0.0003 0.0003 0.0003 0.0003 0.0001 0.00004 0.00003 0.000006 0.2-3 X 10-10 4-8 Table 4.6 Major Categories of Elements in Manganese Nodules for Purposes of this Study. Major and Value Metals Mn, Fe , Co, Ni, Cu, (Zn) , (Mo) (-29% total) Innocuous Nonminor Na , K, Mg, Ca , Al , Ti, Si, P, S, CI (-14% total) Elements Potentially Toxic Ba , La, V, Cr , Ag, Cd , Tl , Pb , As, Sb (-0.5% total) Elements Innocuous Minor B, C, Sc , Sr , Y, Zr , Nb , Ga , Sn, Bi (-0.3% total) Elements Five primary sources were consulted in the selection of those constituents of the manganese which should be flagged as potentially toxic or hazardous, as qualified above. These sources are: 1. "EPA Hazardous Substances, Designation, Removability, Harmful Quantities, and Penalty Rates", Federal Register, December 30, 1975. 2. EPA Primary Drinking Water, Proposed Interim Standards, Federal Register, March 14, 1975. 3. U.S. Public Health Service Drinking Water Standards, 1962 (HEW Publication). 4. N.I. Sax, "Dangerous Properties of Industrial Materials", 4th edition, Van Nostrand Reinhold Company (New York, 1975). 5. Kirk-Othmer, "Encyclopedia of Chemical Technology" 2nd edition, Interscience Publishers (1970). As a first step, cation and anion forms of the elements contained in the nodules and of the major elements in the process raw materials were enumerated. Table 119.5, "EPA Harmful Quantity Categories", of Source 1 was then culled for toxic or hazardous substances which are combinations of the possible cations and anions. The resulting list of some sixty inorganic substances proved to have limited usefulness for the following reasons. In practically no instance would the flagged compounds occur at any stage of nodules treatment as distinguishable and isolatable entities. To the contrary, in nearly all process streams, the toxic components would occur as dilute solute species not specifically associated with other chemical species in the form of the indicated compounds. Further, substances which could be present and which are hazardous by reason of being chemically aggressive (for example, molecular chlorine and its hydration products, strong acid, strong caustic, etc.) would be introduced into the process as materials other than nodules. Finally, the derived list of toxic and harmful substances, although long, was not comprehensive. It did, however, include most of the inherently toxic elements, such as arsenic, cadmium, lead, etc. that may be present in the nodules in nonnegligible amounts. A guide to the degree of toxicity of selected nodules constituents was obtained by reference to the EPA Primary Drinking Water Standards. For the inor- ganic materials of concern to this study, the concentrations are represented in Table 4.7. Other than for mercury, not present at a perceptible level in the standard nodules, the lowest permissible concentration is 0.01 mg/1, or approx- imately 0.01 ppm. Barium is considered to be appreciably less toxic (witness the widespread use of BaS04 in medical x-ray examination) than the other cations listed, but its assumed content of 0.32% in the nodules indicated that it be flagged as potentially toxic. 4-9 Table 4.7 Drinking Water Standards for Inorganic Substances Present in "Standard" Nodules / / / / Limits Substance mg/1 Arsenic 0.05 Barium 1.0 Cadmium 0.010 Chromium (VI) 0.05 Cyanide 0.2 Fluoride 0.6-1 Lead 0.05 Mercury 0.002 Nitrate (as N) 10. Selenium 0.01 Silver 0.05 *USPHS Standard, 1962. Federal Register, March 14, 1975 There are other nodules constituents known or suspected to be toxic, for example, thallium. A check of Sax's Handbook and the Kirk-Othmer Encyclopedia confirmed this classification for thallium and suggested the inclusions of lanthanum (not highly toxic, but 0.13% of the nodules), vanadium (0.032%) and antimony (0.0024%). This completes the roster of ten potentially toxic nodules constituents. In addition, cognizance was maintained, in considering process stream and waste compositions, that there exist recommended concentration limits (USFHS) in drinking water for several of the major and value elements, as follows (concentrations in mg/1): Mn (0.05), Fe (0.3), Cu (1.0), Zn (5.0). At the indicated low levels, the presence of these elements is objectionable on primarily esthetic or practical grounds. 4.2.3 Hazardous Constituents of Process Supplies It is recognized that nodules processing will entail substantial quantities of raw materials, some of which are known to contain potentially toxic and hazard- ous constituents. However, it will be impossible to quantify the amounts of the flagged elements introduced into the process by way of raw materials until it becomes possible to specify the detailed composition of the coal, limestone, ker- osene, etc. This specification would require identification of a source for each raw material, the source, in turn, being determined to some degree by the site selected for the onshore processing plant. An indication of why it was not considered feasible to include contributions from raw materials in the scrutiny of toxic and hazardous elements involved in nodules processing is given by the tabulated analyses in Table 4.8. For some minor and trace elements in this composite coal analysis, the standard deviation actually exceeds the mean value, an indication of extreme variability. Variability of composition is also the rule for other major raw materials such as limestone. Table 4.9 displays an analysis for a specific limestone, which confirms that this seemingly innocuous material can also contain small quantities of potentially toxic and hazardous elements. However, as with the manganese nodules, toxicity and hazardous nature do not develop unless the flagged elements become concentrated in certain chemical forms. 4-10 i Table 4.8 Mean Analytical Values for Chemical Elements in Coal Source: Illinois State Geological Survey Constituent Mean, ppm Std. Dev. Constituent Mean, ppm Std. Dev. As 14.02 17.70 Pb 34.78 43.69 B 102.21 54.65 Sb 1.26 1.32 Be 1.61 0.82 Se 2.08 1.10 Br 15.42 5.92 Sn 4.79 6.15 Cd 2.52 7.60 V 32.71 12.03 Co 9.57 7.26 Zn 272.29 694.29 Cr 13.75 7.26 Zr 72.46 57.78 Cu 15.16 8.12 Al 12900 4500 F 60.94 20.99 Ca 7000 5500 Ga 3.12 1.06 CI 1400 1400 Ge 6.59 6.71 Fe 19200 7900 Hg 0.20 0.20 K 1600 600 Mn 49.40 40.15 Mg 500 400 Mo 7.54 5.96 Na 500 400 Ni 21.07 12.35 Si 24900 8000 P 71.10 72.81 Ti 700 200 Moisture 9.05% 5.05% Table Ash 4.9 11.44% 2.89% Trace Elements in a Missouri Limestone Source: Boynton, Chemistry and Technology of Lime and Limestone Elements ppm Antimony ND Arsenic ND Barium 620 Beryllium ND Bismuth ND Boron 27 Cadmium ND Chromium 160 Cobalt ND Copper 4 Gallium ND Elements ppm Lead 3 Lithium ND Manganese 108 Molybdenum 5 Nickel 3 Silver 0.3 Strontium 1300 Tin 0.2 Titanium 2700 Vanadium 17 Zinc 32 In attempting to track the flow of trace elements through the process routes, it was necessary to make estimates of distributions at virtually all process junc- tures. The convention was adopted of not specifying greater precision in percent- ages entering a given stream than 0, 10, 50, 90 or 100%. In a great many cases, the probable distribution was not known to even that degree of precision, and unsubstantiated estimates had to be made. This is due to the paucity of information in the literature on the distribution of other than the major and value metals, combined with the low concentration levels of the flagged impurities and, often, uncertain chemical behavior of the impurities in the process media. Although the exercise was performed comprehensively, data on the distribution of potentially toxic elements are reported only for those process junctures preceding a plant output, whether product, waste or other effluent. All distributions are reported for the reduction/ ammoniacal leach process, however, to show methodology. 4-11 4.2.4 Processing Rate Capital and operating cost estimates and evaluations of market impacts summarized in Chapter 3 indicate that three-metal plants will be built with capacities of the order of three million tons per year dry nodules, while less than one million dry tons per year will be processed in four-metal plants. For the nodules compositions assumed (Table 4.1) the revenues required for 10% return for a three-metal plant for the low range case (Table 3.8) would be met by processing about 2.5 million tons per year of water-free nodules. About 0.75 million tons per year would be required for a four -metal plant on the same basis, and wet nodules processing rates would be two-thirds higher in either case. Accordingly, the material and energy balances developed here are based on processing 12,500 and 3,750 tons/day of as-received nodules for three- and four-metal processes, respectively. The anticipated plant stream factor is approx- imately 90%, with full production on 330 days per year. 4.2.5 Product Forms It has been assumed that copper and nickel will be electrowon from solution to produce a product of cathode specification. Since none of the leach solutions will be pure enough to permit direct electrowinning , metals separation schemes are required to produce electrolytes of acceptable purity. It has been assumed that these schemes will be based on the use of liquid ion exchange/chelating reagents, since these agents are currently in commercial use in nonferrous metal production. Cobalt will be produced as a hydrogen-reduced, briquetted powder. Minor amounts of nickel will also be produced in this form. Minor amounts of mixed impure zinc and copper sulfides will be produced as a by-product of the cobalt recovery operations. This will be sold to smelters for further processing. Manganese will be produced, in the four-metal process routes, as either f erromanganese or electrolytic manganese metal. A possibility exists that man- ganese could be recovered from leach tailings of the three-metal process routes. This option will not be considered here, since no data have been published on the amenability of leach tailings from these processes to upgrading for manganese recovery. 4.2.6 By-products The process routes have been specified to minimize by-product formation and maximize the re-use or recycle of lixiviants. While this could require the purchase of additional process materials and the consumption of additional energy, it eliminates the need to establish a marketing operation other than one devoted to nonferrous metals. This applies particularly to ammonium sulfate production. 4.2.7 Energy Sources It has been assumed that the primary energy source for the nodules processing plants will be coal. Since the energy requirements are quite high, access will be required to rail systems for the delivery of the required amount of coal, and provisions will be made at the plant site for its storage and reclamation. The coal composition assumed for this study is summarized in Table 4.10. Since sub- stitution of heavy fuel oil for coal would be possible, however, its consumption has also been calculated. 4-12 Petroleum products will be required for use only by plant vehicles and for minor process uses for which the consumption of coal is not a convenient way to deliver energy. Fuel oil or gas would be used only for those process applications for which a compelling economic and processing advantage can be shown vis-a-vis the use of coal. Table 4.10 Assumed Coal Composition Type: High Volatile Bituminous C Analysis: 65% Carbon 6% Hydrogen 18% Oxygen 1% Sulfur 1% Nitrogen 9% Inerts Higher Heating Value: 11,400 BTU/lb The metals recovery sections of the nodules processing plants, in particular, are energy intensive and high consumers of power, while other operations demand large quantities of process steam at intermediate to low pressures. It has been assumed that steam will be raised at high pressure and expanded through back- pressure turbines to provide for the delivery of both process steam and some pro- cess power. .The balance of process power requirements would be met by purchases from local utilities in order to minimize the capital investment in plant facili- ties which are not directly required in the production of metals from nodules. 4.2.8 Water It is assumed that an ample supply of water for process use is available from a nearby source. Water with a hardness of 250 parts per million calcium carbonate equivalent and carrying 200 parts per million suspended solids has been assumed. Most of the consumption is required for evaporation in cooling towers and to carry leached tails and other process solid wastes as slurry to disposal. 4.2.9 Materials and Supplies The consumption of process materials and supplies directly used in nodules processing has been estimated in developing the material balances. The composi- tion of the major process materials is summarized in Table 4.11. No attempt has been made to calculate the consumption of maintenance materials and replacement parts such as furnace refractories, new anodes, etc., but the value of these items typically runs to about 1-2% of the capital cost of the plant annually. With the exception of the reduction/hydrochloric acid leach process, the plants are assumed to be self-sustaining, requiring only make-up of lixiviant and supplies consumed in processing. For the hydrochlorination route, it has been assumed that a buyback arrangement exists for the chlorine evolved as hydrogen chloride is consumed in the process. 4-13 Table 4.11 Composition of Process Materials and Supplies Gases Ammonia Hydrogen Hydrogen Sulfide Chlorine Nitrogen Liquids Organic Sulfuric Acid - Nitric Acid Sodium Hydroxide Fuel Commercial Anhydrous Commercial 99% Minimum (liquid) Commercial 97.5% (liquid) Commercial 99.5% Commercial Liquid ion exchange/chelating agent dissolved in diluent at concentration appropriate to each process. Commercial 93% Commercial 60% Commercial 50% Vehicular and combustion fuel as required Solids Ca Mg Limestone 80 (CaC03) 15 (MgC03) % Inerts Lime 79 (CaO) 12 (MgO) 9 Flocculants Additives Sodium Sulfate Boric Acid Carbon Commercial Polyelectrolytes Commercial Electrov/inning Additives Anhydrous, Photo Grade Commercial Granular, 99.9% Commercial Activated Carbon Water Treatment Chemicals Borax - Technical, Anhydrous 99% Electrode Paste Salt - Commercial Rock Salt 4.2.10 Process Waste and Emission Treatment Procedures It is assumed that the design of any nodules processing plant will be suffic- ient to ensure compliance with applicable federal and local regulations concerning the discharge of solid, liquid and gaseous effluents. For the present study, we have assumed that in the base case, this will require that: • All combustion gases will be scrubbed with limestone slurry for sulfur removal and that the combustion processes will be controlled so as to permit compliance with nitrogen oxide and sulfur oxide emission regu- lations . • Process wastes which are combustible will be burned on site. • Adequate measures will be taken for dust control at appropriate places within the process, with effluents discharged after further treatment. • Fugitive emission control in high temperature operations will be achieved by the use of hooding and high volume ventilation, with the fugitive gases being scrubbed prior to release to the atmosphere, as required. • All vents on process tankage will be manifolded to scrubbers or protected with conservation units. 4-14 • Process solid and liquid waste, including plant run-off, will be combined with leach tailings or granulated slags, neutralized if required, and piped to lined tailings ponds for containment. The impact of alternative methods of solid and liquid waste treatment is ad- dressed in Chapter 11 for all process routes. 4.2.11 Plant Waste Products * The outputs from each of the processes, other than the metal outputs produced for sale, can be classified as waste products. The waste products, or streams, will be in the form of gaseous emissions, solids, liquids, or a combination of the latter two (e.g., slurries, slimes). Gaseous emissions will be treated, as neces- sary, and this treatment will result in additional liquid and solid waste streams. The waste streams listed are the net outputs from a given process after feasible treatment and/or recycling within the process itself. The majority of the solid wastes will be derived from the nodules themselves, the lime used in the process, and the coal consumed. The waste outputs and the physical and chemical constituents of the major components thereof have been defined in the process descriptions and in the process flow sheets. The poten- tially toxic elements (ten elements comprising approximately 0.5% of the assumed as-received nodule composition) have been identified and tracked through each step of the process. For each of the five processes, summary tables have been developed which list process waste streams, the total flow in tons per day, and the expected method of disposal. The toxic elements expected to be included in a waste stream are also listed. Variable amounts of potentially toxic elements will exit the proces- ses as impurities in the saleable metal products and are, therefore, of no further concern. 4.2.12 Potential Accidents with Pollutant Release The nodule reduction/refining processes utilize some toxic and hazardous mat- erials which are considered environmentally detrimental: mineral acids, ammonia, kerosene, hydrogen sulfide, etc. En-vironmental hazard analysis focuses upon accidents which have the potential of releasing such materials into the environment. Indus- trial accidents which might not release pollutants, other than ordinary smoke (inert particulates and hydrocarbon gases) , uncontrolled into the surrounding environment are considered outside the scope of this hazard analysis. Primarily, the accidents of concern would involve large releases of liquids and gases - a restricted subset of industrial accidents. For this reason, industrial accident statistics are not completely applicable for estimating the potential occurrence of pollutant releases, but do serve as a guideline envelope to those accident situations which could result in releases. Basic environmentally protective design of the nodule plant is assumed to a reasonable level, including avoidance of unusual foundation problems, 100-year flood zones, active faults, etc. Containment of spills from stored hazardous and toxic reagents by diking, berms, and sumps, as required by Federal regulations for safeguarding water quality, would be incorporated, and a contingency plan for 4-15 controlling and containing spills is assumed to be in effect, along with personnel training. Avoidance of elevation and slope exposures for large volume items, which could create high velocity flows in the event of rupture, is an example of a simple design criterion which would reduce the impact of large liquid spills. An emer- gency standby sump at the site to provide a two-day process surge for tailings pipeline shutdown is assumed to provide spill protection for the site. The liquid and solid materials which could be released are not exotic organic substances of high toxicity and persistance, such as pesticides. Thus, with the exception of fragile local ecosystems in the drainage path and watersheds sup- plying potable water, potential effects of releases would be short term in nature. The toxic gases which could be released in a system rupture could spread to con- siderable downwind distances with seriously high concentrations in the event of large volume releases under stable atmospheric conditions (adverse for dispersion). The hazard is greater from storage accumulations than from the plant processes because of the volumes involved. The analysis was limited to activities within the plant boundaries and therefore did not consider transportation hazards. Further analyses could compare the risks posed by relatively large storage and infrequent deliveries with risks posed by smaller storage capacities and more frequent deliveries. The primary mitigating factors for the gaseous hazards is assumed to be avoidance of sites with adverse dispersion characteristics or, alternatively, avoidance of populated zones, protection of storage vessels from sources of rupture, and assurance of storage vessel reliability. 4-16 5.0 REDUCTION/AMMONIACAL LEACH PROCESS 5.1 PROCESS DESCRIPTION 5.1.1 Summary Process Description Copper, nickel, and cobalt can be recovered from nodules by a reduction/ ammoniacal leach process. A simplified block diagram of the process is shown in Figure 5.1. The first step in this process is a high temperature (625°C) reduction of manganese dioxide, the major mineral in nodules, to manganese oxide by producer gas, which is rich in carbon monoxide. The effect of this reduction is to disrupt the mineral structure and release the contained metals. The metals are removed from the reduced nodules by dissolution into a strong aqueous solution of ammonia (10%) and carbon dioxide (5%), at low temperature (40°C) and atmospheric pressure. The metal-bearing solution decanted from the nodules is subjected to a series of purification steps, in which it is contacted with an organic medium which selec- tively and separately removes the copper and nickel from the aqueous solution. The metal values are, in turn, selectively removeu from the organic fluid, and transferred to acid aqueous solutions, which accumulate copper sulfate and nickel sulfate. The metal products, cathode copper and nickel, are produced from these solutions by electrodeposition . Cobalt is then recovered from the aqueous ammonia /carbon dioxide solution by contacting it with hydrogen sulfide, which precipitates the insoluble sulfides of cobalt as well as small amounts of copper, nickel, zinc, and other metals not removed in previous steps. The solids are removed from the aqueous ammonia/carbon dioxide solution and contacted with air and hot (100°C) sulfuric acid to selectively redissolve the cobalt and the small amount of nickel present. The undissolved sulfides are sold as minor products, and the cobalt and nickel are recovered from solution in powder form by selective reduction with hydrogen at high pressure (34 atm pressure) and temperature (185°C). The reduced nodules residue, from which the major portion (98%) of the soluble metals has been removed, is contacted with steam (at 120°C, 2 atm pressure) to remove residual ammonia and carbon dioxide. The ammonia/carbon dioxide/steam mix- ture is condensed and, together with the aqueous ammonia/carbon dioxide mixture from which cobalt has been removed, is recycled to extract more metal values from freshly reduced nodules. The steam-stripped nodule residues are combined with smaller amounts of other process solid and liquid wastes and sent to containment. Plant services include facilities for generating the producer gas used in nodule reduction, raising the necessary steam and part of the power required for process use, supplying the make-up water and cooling required, and providing for materials handling for process materials and supplies. The high temperature reduction of nodules with simulated producer gases and subsequent extraction of metals with ammonia/carbon dioxide solutions has been extensively studied and reported. This approach is basically the same as is cur- rently used in recovering nickel from laterites by the Caron process, and can be 5-1 t O li) XT- - - J o TV z —I o ui o o a: 0- gO 5 I CO 6 o g M 1 i3 Z J 2 o ^, I- s . t i z t s |n J ^- ^ TT II I 5-2 considered a variation of currently available technology. The metal separation and purification scheme, however, is specific to nodules and is complicated by the chemical similarity of copper, nickel, and cobalt. Separation and purifica- tion of copper and nickel by selective extraction with organic compounds (liquid ion exchange reagents) is currently practiced in the extractive metallurgy of copper and nickel. However, in these cases the aqueous solutions contain primarily one metal, the others being treated as impurities, not products. Cobalt recovery from precipitated mixtures of nickel and cobalt sulfides derived from laterites is also currently practiced. The details of the procedures used to purify the leach solutions prior to reduction, however, would differ somewhat from those used for nodules because of the differences in amount and content of impurities. The generation of producer gases from coal or oil for the reduction of nodules and all other plant services represent the utilization of known technology essen- tially without adaptation. The basic process configuration proposed, then, is a modification of the well established Caron process for recovering nickel from laterites. The major differ- ence is in the metals separation and purification steps, and represents an exten- sion of elements of existing technology to a new application in nodules processing. The detailed design bases used in developing the process description are summarized in the criteria sheets in Section 5.7 A major alternative to the proposed process configuration involves the low temperature, low pressure aqueous reduction of the nodules using high strength car- bon monoxide gas. This option will be investigated in detail in another section of this report. Other variations in the process are possible, particularly with respect to the specific metals separation and purification schemes used. However, they would likely have only minor impact on overall process inputs and outputs. 5.1.2 Detailed Process Description The reduction ammoniacal leach process is a three-metal process in which copper, nickel and cobalt are liberated by an oxidizing ammoniacal/ammonium car- bonate leach following high temperature reduction of manganese dioxide by a syn- thesis gas. Copper and nickel are coextracted by liquid ion exchange reagents and selectively stripped and recovered as electrowon cathodes. Cobalt is recovered from the raffinate by precipitation with hydrogen sulfide and is recovered from the sulfide precipitate by selective leaching and hydrogen reduction along with some nickel, zinc and copper. The metal-free raffinate is recycled to provide leach liquor and for washing tailings from the process. Ammonia and ammonium carbonate are recovered from leach tailings by steam stripping. Detailed descriptions of each segment of the process are given in the following flow-sheets. (Dwg. NC-OOOl-Am, Fig. 5.2) Wet nodules are reclaimed from storage and fed through a primary cage mill where they are reduced to -7/8". They then pass to a fluid bed dryer where surface and pore water is removed at 175°C by direct contact drying with combustion gases. Bed overflow is reduced to -200 mesh in a secondary cage mill, and dryer and mill fines are removed from off-gases with cyclones and an electrostatic precipitator. Off-gasses pass to gas treatment for scrubbing, while dried nodules are trans- ferred hot in an enclosed conveyor to reduction. (Dwg. NC-0002-AM, Fig. 5.3) Dried nodules are reduced at 625° C in a two-stage fluid bed reduction roaster by contact with producer gas derived from coal gasification. At re- 5-3 o V) »- vu ui 5-4 "FT mi m 'i Ul 6 r, u 3 (0 1 f^ i 1 -^^ I 1 — /" V 1 "'"1 A "'— /' : \ -~ _ — . —/ I I -I. ■^ A ^^ :2 I :2 tii^i.-;:^ I s 5 ^ HI 2 < « o! o , o i :2 ; ■i__/ O H |i4 :S ^ «; i JU •i 00 ! ' ii^ »""- 3 V \ pi W Q i^ H — I 5-J Q Q N a- a. « 1.1 4- F <^ •^-r-- -' n ^3 H 1-, _1_L -s ^-' ?. 1 ^ -S) a -j- : V" i -.^ .) • IVJ ;_ .- ■"! 'J '4 < in J "' VJ Ul ^■I :;: ' '';^ a u L - - - 1 1 * ; /' 6 - — L [ — I m 1 t- O >t - vu n oj -i ^ -" a: vS ■<} ^ 1 s s = 4 s s ^ K "M ^'■§ - 2 ? s § = s s ^ - S f! 5 ^ - 3 . I ^ < ■ft - 1 g' .^Jl_:l^^J * : o I -3;^ IS 5 ! o O I o 1 O M 5-9 5-10 from copper electrowinnlng . The copper electrolyte is also heated/cooled on passage from/to stripping to permit operation of electrowinning at a higher temp- erature. Make-up organic is added to the stripped reagent to offset degradation and soluble organic losses to pregnant liquor, and the metal-free organic is recycled to extraction. (Dwg. NC-0Q07-AM, Fig. 5.8) Cathode copper is recovered from the strong electrolyte from liquid ion exchange using conventional technology. Starter sheets are deposited on titanium blanks from strong electrolyte in the stripper section and are removed, washed, looped, and returned to the commercial section as starters. Full-term cathodes produced in commercial section are washed, unloaded, and prepared for shipment as cathode to sale. The greater portion of the weak electrolyte is recycled to the liquid ion exchange section for stripping but, since a small amount of nickel is co-stripped with the copper and not deposited in electrowinning, it must be purged from the system. The purged electrolyte passes through purification cells where copper is removed by electrowinning to depletion. The decopperized electro- lyte passes to vacuum evaporators where water is removed and nickel sulfate pre- cipitated from the resulting highly acidic solution. The nickel sulfate is removed and sent to cobalt recovery, while the acid is returned to the process where, with make-up acid, it is used to redissolve scrap copper for return to the commercial cells for deposition. Sufficient steam, wash water, and make-up water are added to the circuit to offset water which is vaporized and carried, with evolved oxygen, from the electrowinning cells. (Dwg. NC-0008-AM, Fig. 5. 9) Nickel is recovered from the strong electrolyte by electrowinning in a manner similar to that used for copper recovery. In nickel electrowinning, however, cathode bags are used, and sodium sulfate and boric acid are added to the electro- lyte to control its conductivity and pH. Dissolved organic carried from the liquid ion exchange step is removed by absorption on activated carbon prior to any electro- lytes passing to electrowinning. Also, the starter sheets are pickled in sulfuric acid prior to use in the commercial cells, and nickel scrap is redissolved in ammonia-containing raffinate and recycled to the pregnant liquor, rather than to the recycled or make-up acid solution. The electrolyte purge required to remove impurities from the electrowinning circuit passes to a sulfide precipitation reactor and then to cobalt recovery for retrieval of metal values. (Dwg. NC-0009-AM, Fig. 5.10 ) Cobalt, along with unextracted copper, nickel, and zinc is recovered from the raffinate from liquid ion exchange by precipitation with ammonium hydrosulf ide, which is produced by sparging hydrogen sulfide into an excess of ammonia solution. The sulfide precipitate is separated from the raffinate in a clarifier, with the clarifier overflow being recycled for tailings washing and the underflow passing to stripping for ammonia recovery. The ammonia-free sulfide slurry is mixed with purges from copper and nickel electrowinning and cobalt, recovered from stripping the liquid ion exchange reagent. The mixture is pressure leached with air to per- ferentially dissolve the nickel and cobalt sulfides, leaving copper and zinc sulfides undissolved in the residues. The latter are removed by filtration and sold, as minor products, to smelters for recovery of metal value. Following pH adjustment and reprecipitation with hydrogen sulfide for final removal of any zinc and copper solubilized in the first leach, the nickel/cobalt sulfate solution is heated and autoclaved, and nickel is reduced with gaseous hydrogen. Sufficient ammonia 5-11 IT A ■■'■ -a '. f iJ '5i ^. -i-t. ^^-r:: 4: .'^' "i i :2 :2 1- U U) _i tu ft! Q. a . 8 Q U 8 8 < ^ -J Q 01 8! ^ "? ^^ j b v^'' .1 1. ^^^1 - t .>; ♦!_ 1 1 1 - 1 - » ^ 8 hi -:• fT' Wl r I .p.. ,,|'.. L,^'- LL--,. '^^ < * '' ! ^ U\i. 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After densif ication through repeated recycle, the nickel powder is removed, washed, and passed to drying and briquet- ting for sale. The largely nickel-free cobalt sulfate solution passes to an evaporator/crystallizer where the remaining nickel and cobalt are precipitated as the double salts with ammonium, sulfate. Exces;3 ammonium sulfate is purged, and the salts are redissolved in strong ammonia sc>lution. The cobalt in solution is oxidized to the cobaltic state with air. This permits the cobalt to remain in solution when the stream is subsequently acidified to remove nickel salts, which are then separated and recycled to the pH adjustment step. The nickel-free solution is then heated and autoclaved for removal of cobalt by hydrogen reduction. Sufficient ammonia is added to neutralize the acid .generated. The cobalt powder is dried and briquetted for sale, while the ammonium sulfate is purged to ammonia recovery. (Dwg. NC-0010-AM, Fig. 5.11) Tailings from the CCD wash circuit are preheated and stripped for ammonia recovery by counter current contact with steam in stripping towers. The steam and ammonia vapors from stripping are combined with vaport; from other ammonia strippers and pass to an ammonia absorber /condenser tower where most of the ammonia and ammonium carbonate are condensed. Vent gases pass to a carbon dioxide absorber where they are absorbed, along with the required make-up carbon dioxide obtained from boiler off-gases, in make-up scrubbing water. Ammonia is recovered from all process vents by counter current contact with make-up scrubbing water in an ammonia absorber and vent scrubber. The ammonia-free gases pass to stacks for disposal. Ammonia is recovered from the ammonium sulfate purges fnom cobalt recovery and the liquid ion exchange washing step by reaction with slciked lime. Steam is blown into the mixture to strip the evolved ammonia, and the vapor is condensed, with the condensate returned to aqueous ammonia storage for recycle within the process. The gypsum slurry from the lime boil is cooled, along with the slurry from the tailings stripping, and passed to tailings surge. This is; combined with process solid and liquid wastes and plant run-off, treated for pH control, as required, and pumped to the tailings impoundment area for disposal. (Dwg. NC-0011-Am, Fig. 5.12) Facilities are provided within the plant for receiving and reclaiming raw nodules, coal, lime and limestone, ammonia, and other process materials and fuel. (Dwg. NC-0012-AM, Fig. 5.13) Plant services include process and cooling water supply' and treatment, steam raising and power generation, gas treatment, and reducing gas preparation. Make-up water is clarified and softened for distribution to the proct^ss, as required. Additional treatment is required for cooling tovjer water make>-up, boiler feed water make-up, and for supplying plant potable water. Off -gases from manganese reduction are burned, along with additional coal, in the main boilers to raise the required process steam and generate a portion of the powei" required in the process. Following particulate removal, the flue gases are combined with other process off-gases and pass to gas treatment, where sulfur oxidi2s and other acidic constituents are removed by scrubbing with limestone. The scrubbed off-gases are reheated, combined with scrubbed vents from ammonia recovery, and stacked. Gas for nodules reduction is produced in a two-stage entrained- flow gasifier in which coal is mixed with preheated air and high temperature steam for the produc- tion of a co-rich reducing gas. The gasifier product passes directly to reduction, following particulate removal and energy recovery, with sulfur removal taking place after combustion along with other boiler flue gases. 5-15 r '^' « i _l 1 1 i ^ ; 1 1 1 b 1 |. o w P 5-17 FT IV I V) I ^ ft M "¥ ^ l^t ru 2^ I M ■^ i -d -l iH eu CO .H CO CJ C o •H 4-1 TD -4- x o 3 vD M 0) J-i kJ en U CO o- QJ 0) iH a: CO 60 tH c 3 U »N CO O 3 so o o CNI o o o CO <}• fO f O O CM O CM O C/1 CO O CN O O to CO -H rH C C a CJ en < s s ro o o .H CM O o CJ^ O CN O CO •H tH 4J 00 CO <; w x; o CN CN O . 00 O <~M O CN •H tH 4J O CN O U O rsi o cO -H tH 4-1 a en < w CN - en < W ro PC CO CN O o >. O o U o CO c c a 0) tH s s s fn U -i 00 -l (-1 o cu •H >-l p 3 3 tH O CO C 4^ C C M 3 5-23 TO H U) e cti V) cu w u 0) •u o ixi o i-l 4-J Pi c QJ -C 3 u rH n) m (U M-l hJ W rH TS CO •H CJ 3 nj cr •H •H C hJ o M-l 1 o < c c o o •H •H 4-1 ■U •H O Cfl 3 O 13 a 0) e C^ o u u (U U3 ^ >i to J3 i-i 60 3 i-i U-l Sj 3 M-l O rH O 1 CO en < C« 1 •H c >! 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C — c~j CN rv e u > 60 C 3 ,a 60 G ■H ■O 3 .H U X 01 n 3 0) 5-25 5.4 DISTRIBUTION OF SELECTED ELEMENTS IN MANGANESE NODULES PROCESSING: RE- DUCTION/AMMONIACAL LEACH PROCESS 5.4.1 Explanation and Qualifications A metallurgical process may be construed as a network of branching streams, leading from material and energy inputs to products, wastes and other plant effluents. At a process juncture, as we have used the term, an influent stream (or streams), which might consist of solids, a solution, a slurry, etc., becomes partitioned into two or more departing streams. Each departing stream either becomes an influent stream to a subsequent process juncture or is directed out of the plant. The designation of a given stream in a process flow sheet as a departing stream from a juncture is determined by whether the stream under consideration actually leaves that portion of the process. For example, in nickel electro- winning from the strong electrolyte produced by liquid ion exchange, the weak electrolyte (i.e., partially depleted in nickel) that is returned to the mixer/settlers to strip additional nickel from the loaded organic liquid does not constitute a departing stream. In fact, were it not for the weak electro- lyte purge, in this case to cobalt recovery, elements co-stripped with the nickel and not incorporated into the electrowon cathodes would build up to saturation in the electrolyte. The purge from nickel electrowinning, the nickel cathodes and the collected wash solids are departing streams in the context of this study. In considering distributions of selected elements, we have excluded tran- sient changes in composition, as incurred during start-up or upsets, and confined attention to the steady state. In steady-state operation, the com- positions of streams, solutions in surge tanks, etc., have become sensibly constant with time as a consequence of the rates of addition of chemical constituents to a given volume element having become equal to their respective rates of removal. Furthermore, gaseous streams, even when exiting the plant, have not been considered as departing streams for the ten selected elements. Rather, it has been assumed that gaseous effluent streams have been treated to remove the elements in question, which are then returned to the process in the form of solids or slurries. It has been assumed that available technology, including combinations of wet scrubbing and wet and dry electrostatic preci- pitation would be effectively applied in nodules processing plants. 5.4.2 Distributions at Process Junctures In the absence of specific information on the fates of selected elements, it is necessary to resort to estimation based on known or probable chemical behavior or on inference from behavior in analogous processes, or to employ one's judgment. In many instances, the judgment will be little better than an educated guess. The degree of uncertainty in an assignment of a distri- bution should be roughly reflected in the following tables by the indicated Confidence Level and Source/Comments. Where there is no entry under the latter heading, judgment or guess is to be inferred. 5-26 In view of the uncertainties involved, the partitioning of the selected elements has been drawn no more finely than is indicated by the allowed per- centages: 0, 10, 50, 90, and 100. That is, a given element is assumed not to enter stream A; a small fraction does; it splits roughly equally between streams A and B; most of the element reports to stream A; substantially all of it does. In cases of more than two departing streams, distributions are re- ported in multiples of 5 percentage points, but the real uncertainty is greater, not less. It is recognized that the cumulative effect of errors in assignment can be very substantial — even orders of magnitude. However, the framework is herein provided for improving the accuracy by supplying hard in- formation when the data become available. The following table (Table 5.6) of toxic element distributions and stream concentrations constitutes a complete working set for the reduction/ammoniacal leach process. It is included in this section to illustrate the methodology for predicting the manner in which selected elements will bo apportioned among process streams. In the four following sections, which contain descriptions of other likely processing routes, the detailed distribution tables have been omitted, the essential information being displayed in the form of attenuation factor and waste composition tables. 5.4.3 Diagrams of Attenuation Factors As emphasized earlier, in evaluating toxicity and hazard potential, con- centrations and chemical forms rather than total flows are the significant measures. In the absence of chemical analyses, steady-state concentrations can be estimated from a comparison of the flow of a selected element with the total flow of the given stream. The level may or may not be significant, depending upon the specific chemical species involved. We have described element flows by the iterative application of assumed distributions at process junctures. The application must evidently be carried out sequentially, as illustrated in the following diagrams. The running products of assumed fractional distributions are referred to here as "attenua- tion factors" for process streams. By the convention employed, an attenuation factor is a quantity which, when multiplied by the rate of input of the given element, as contained in the nodules fed to the process, gives the flow of that element in the given process stream. This procedure, when carried out for all process junctures using the distributions summarized in Table 5.6, effectively tracks the flow of all toxic elements throughout the process and clearly shows to which streams they ultimately report. Thus, the attentuation factor diagrams for the reduction/ammoniacal leach process. Figure 5.14, shows that all of the Barium in the incoming nodules . (stream 1) ends up in the steam-stripped leached tails (stream 160) . Then, since the incoming nodules flow rate and Barium content are known — 12,500 tons/day and 3040 ppm respectively — and the flow rate of steam-stripped, leached tails is known — 8,180 tons/day — the Barium content of this waste stream can be readily calculated to be 4,645 ppm. Similarly the production of cathode nickel, 106.8 tons/day, amounts to .85% of the incoming nodules weight, and since 11% of the nodule's Silver content reports to the cathodes, it is 5-27 (0 H 00 C 3 o X a, T3 a) w Pi CO o H CO o -*< H 5 S3 Pi; < H W CO i-J QJ 1-1 c > 0) O) n) P. 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P 4-1 -H C t-i O W O PM CN 4J -H O O O o O O o O o pi !z; (J • H M b r-- tn 4-) 5~s ^ rH o C7\ r-i t3N CT^ o> •H 3 rH CO M cn ^-' P 43 vO Pi rH CM Pu P o <^ C <3S c Z pL4 ^>— ' •— N O o •5 c ■H - , •H Z >H -H 4-1 CN a-) pd o pi s tt) ^ O U M w ~^ 4-1 M VC o O CN -;r o o O O o C <: H > .-H C 4-1 (^ o t3^ - cn pi M CM C O H -- OJ ;:3 u H M CJ CO 1— o o u w W ^-q + CO Pi CJ <; ;j ^ Cl, < CO w e 5-35 CO (0 0) y o u Ph ^ u rt -3 l-l (vl o rt •H c o R S < ^ ^^^ -o u 0) o 3 •H C i-t •H o U 3 c -a o aj u a; ^£) in dJ rH J3 a H "F" ^ i ■/. I ■£ i f ^ i . , I* : tH •^ u 1 U ■ * . 4 'J 1^ V4 01 Vi 1 c ■ c CO to ■->-. 'J I I : v ■ 3 !^ ^ cTii o ^ -J ^ =- I 'J -J won 1 < I IJ o w o I -u — I O «N !-H 3 i H -'jC ^ I I L H < O u Q z < ►J tJ u z o o i-l o -a -l H IJ 3 •H CO XI x; CU XI X ij >i -a >> O u J2 ^ •H j= M t3 I -* cn O S o 4J 0) >• + CO •H O >-l •H J- 1 -O 1 >-l .H CU 1 TS l-i ■H T3 CU T3 >^ (U Q ^ C 60 CO C 4-1 4J O CO •H X ^ 4-1 4-J C ^ O J2 cu x: tn C 3 in u D. CU u iH 0. .H + w e 'H -a ^ B o cn CO J3 >% 3 CO 0) cn C 3 XI -H •H C CO .H O CO W 1 ■H g •H O CO ■ cu CO •H -a o 60 3 cu CO CJ ^ 1 M H P4 1 1 M-l CO o CO •H 02 1 V4 u u )-l S-4 M n) •H a) 1— 1 •H o o •H •H P3 14-1 o aj CO O o CO CO c c > O CD OJ 1 ta Ph a< 1 &M fa P-. 1 1 U T3 kJ 1 c c o C/l 1 c iH U Q O . ct) M OJ -H E iH o o O (N a. hJ ^^ O 4J Q. 1 o vO 1 o CN 1 1 -. C (0 p. -* CN r~.' Q ro CO tij o u O 4J rH rH c ■H hJ -^ tH M w cn „ O H C 1 C o S CQ c/o O -i o O c O O •H w w S 4-1 -H C7^ CO in Ov iH in iJ 1^ S • ■H " •H oi 4-1 <■ 4-) w ^ CO f^ (J Di > C 4-1 U c 3 O O -H C 4-1 O o in <■ CO fc o e 3 0) 5 o rsl cn vD H -— ) w ^ LO CO C3^ J ai rH .HOW 1 cn 1 ^ 1 1 en M CO u-i c ai tn o <: M c O H OJ « M M U CO o Z cn w o • s § -* o 1 Ph I-H S ro 1 4-1 1 J < ^-^ u < H ' P- < m w e 5-37 CO ., O Cfl CO u •H O C -U o a CU CO i CO u CO 6 M-i CO 0) CO o ^-- C Cfl o c o s ^4 O -H o a •H 4-1 ^-1 1 ■U CO u XJ O 3 <: 3 3 C CO TJ OJ (U -l-J • o U U 4-1 4J CO c OJ CO !-i x: T) rH CU 4-1 0) D. 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CX CU O e >, a, (A) •H •H U3 •H X U OJ 4-1 4J • 4-J O CO •H CO C/3 O rH S CU •H >-l (U >-l Q a S-1 -a 3 0) T3 x: #> ^ CU o • CU a CO CU X o CU ■ CO u iH •iH <-{ ft s to <: in un o r-l CO in o o > en o u en o ^ en U < rH Eh XI fl4 O cn o IT) o r-H o en o H rO > tn <: O o O rH o en H H pq M P(< H OT o o o T! to X! rH Xi U << M H fl4 rH O o O o o (rt H LO cn rH cn to ■ • • o 04 to •H Q (0 (rt M fT> CQ P-J > CJ rtl C) H o o o o o o en en cn H 5-38 s o w M g w H w g M Pi <: M Ph c; « w Q u (ii M ^ Ph W <; hj Pi CO n rH CO o o w ja rH .n < w H PLI 1-1 in O ro () X3 (\) cC Vj Ti U U U O CO rH n O in "^ 00 00 'H o o o o ^ Xi CNJ O 1^ (Ti ro r~- r~ in o o w XI -H ^ < CO Eh Ph r-H O CO 00 CO in oj o i^ o U 60 T3 > U <: u ■^ o in (Ti o O 1.0 in o O iX> r- Q O o w pei o ^ to > u U ^ c in H in "^ CO in in o o o '3' "^r CO XI tH < W H O >i ■P O u CTi 1— 1 Cl) >i oo O 00 > M OJ O O n iH iH • ■ • u (0 IT) QJ D CO Pi C ro a; en cv) U E > •H hJ > u 2 w U o O Ti o in < en o H Pm in CTi rH O •sf O CO o o <; u 1— 1 IT) o o m r-l o o CNJ ^ M X2 iH rO <; w H PM o o o o in r-l CTi in o o o o (-1 &■} TJ > o < o o o iH iH in in o o ^ "^ o c o •H EH -H> X! (C 0^ a) ^ T) in ^3 G CM !h •H o U U • " H to < zn r-l H P-i o o O O in o o O O O O in o o > 60 T) o o o o in o o in o u Cn TJ > u < U iH in 00 ■^ -H O o o ^ o o rsi O o 00 p w ;=i s 1— 1 H 2 O CJ o ^ — ' s M -;t p:^ PM .H W p., • P- M in PL, pi; O H w U M oi ;:d o M Pm CN y 5-40 0) M >, u u ■H (1) 2 > H \ o -P u rH CI) (fl ^ X! O U 01 Xi rH X! <: W tn cu in rH CTl in o O O o o O O o o O O o U aCTS > u <: u o o in in o o m 4 o o o o o m Xi rH ,n < w H PM in o n •^ IX) VO CN o iH o o o O o m Xt rH ^ < Ui B Ph LD a\ <-\ O ^ o m on O o o o O o o o u 00 -d > O iH u > U in Cn T3 rH in Cn 00 rn <; U CO ^ a: rH o •^ O o ■^ o O o ^ O o o o o o iH ro O o <-{ o -p >i w rH u !h (1) (rt Q) 0) Ti x> > a, o a ^ u U 4-> (U u (Tt o 0^ c; H U) Xi rH <; w EH X) in CTi iH '3' o CO CO o o O iH o o O O o o O O M > U en U H ro < ■^ 04 o O O O CN O o rH O O o o o -Q U) CO <-i X) < rsj Eh ft >sO r- in n ro o KO CNI O o o o O o o o Q H 2 O u 5-41 I Xl r-i Xi TJ r-i ,n U H Ph U H P-i QJ O o 4J o 0-) (U ro (/) 00 O 4-1 r^ r^ O tfl o --I o M v£) u~i o IjS o u-l o ^ O O O o w CO ^-1 O w U > u 03 U T) w kJ > u •H o TS 3 in rH •H O ~;r -H cr o rH4HS. Conversion of the sulf ided metals to powders of Ni, Co, ZnS, and CuS. Criteria Cone, of NH3 Solution: 12.5 wt % Produce a 30% NH4HS solution Temperature rise of 60°F necessitates cooling 10% excess of S Use steam for 10°C increase of stream to 50°C Metal Precipitation : 99.9% Zn, Cu 99 % Ni 98 % Co Temperature: 80°C Ammonium Hydrosulf ide: 33% excess Clarifier Specification : Residence time: 2 hr Clarifying rate: 20 £/min/m2 Underflow density: 5% solids Overflow entrainment: 100 ppm solid Filter Specification: removal of solid to 10 ppm Wash Ratio: 2:1 Centrifuge Specification : a) solids density of centrifuge: 16.6% b) require 50 hp centrifuge c) operate at 2000 g's Steam strip of NH^ to 10%/C02 to 5% Source Confidence Assumed Assumed Assumed Assumed Assumed Co recovery practice in laterite Ni ore processing. Assumed Assumed Assumed Assumed Assumed 5-70 Good Good Good Good Good Fair Fair Remarks Fair Fair Fair Poor T Plant Section: ^o^ialt Recovery (continued) (continued) Criteria Source Confidence Results Leaching Specification: a) Slurry at 40% solids b) 70% sulfuric acid solution c) 1% purge of leach solution d) Air at 20% utilization, 100% converstion to SO4 React Stoichiometric H2S with purge from leaching. Nickel Reduction at 175°C and 500 psig Flash to 100°C Nickel Recovery: 90% Evaporation/Crystallizer - remove 70% of water Use 25% NH3 solution at 100% excess prior to oxidation Cobalt oxidation at 100°C and 150 psig Cobalt Reduction at 175°C and 500 psig Assumed Assumed Assumed Assumed Assumed Assumed Fair Fair Fair Fair Fair Fair 5-71 Plant Section: Anunonia Recovery Operation: Tailings Stripping (Raffinate Stripping) Function: Description: Strip NH3 and CO2 from wash liquor retained with tailings. Criteria Source Confidence Remarks ?':r iuriinj: Conriitions : Pressure: 1.5 aCTn Stripping ratio: 1.2 x minimum Recovery: 99% of NK3 , CO2 Number of stages: Vapor VelQcityi 2 meters/sec. Fefed conditioning: none Energy recovery: by vapor condensation, tailings stripper / feed/bottoms exchange raffinate. sti xpper . Holding Tankage: Retention time: 2 hr Temperature: 40°C Indirect Condensation : Condense NH3/CO2/H2O in shell and tube condenses cooling to 40°C at 1.3 atra. Direct Condensation : Condense vent gases from indirect condenser in packed tower by direct contact with cooled ammonia car- bonate solution at 40°C, 1.3 atm. Vapor velocity: 2 m/sec Number of stages: 1 CO2 Absorber Absorb CO^, from power plant stack gas "slip stream, into NH3 solutions to supply make up. Izsei on .iicirc : practice Standard practice Based on Nicaro practice Poor '■JiiLr.^ tenaencies tailings unknotJn. Good Good Based on Nicaro practice 5-72 Good r Plant Section: (continued) Ammonia Recovery, condensing/scrubbing Cr iter ia Results Absorber temperature: 40°C Absorber pressure: 1.2 atm Absorber efficiency: 99% CO2 (including NH3 absorbers and vent scrubber) Number of stages: 1 Vapor velocity: 2 m/sec NH3 Absorber: Absorb and condense NH3 (and residual CO2) from process vents by direct contact into cooled ammonium carbo- nate solutions. Absorber temperature: 35°C Absorber pressure: 1.2 atm Vapor velocity: 2 m/sec No. of stages: 1 Vent Scrubber : Absorb NH3 from process vents into make up scrubbing water. Absorber temperature: 30°C Absorber pressure: 1.1 atm Exit gas NH3 content No. of stages: Vapor velocity: 2 m/sec Based on Nicaro practice Good Based on Nicaro practice. Good 5-73 Plane Section: Ec-iip-ent Iter.: Function: Description : Material Handl ing Assorted Tanka g e and S tora ge Areas Provide Surge Capacity for Plant Operati "•5 Require:nents Process -aterials and supp lies are recei ved fron various sources , held in stora ge, and retrieved as required. Criteria Source Confidence Remarks Nodules storage: Nodules received as 30% slurry and held wet in con- tainment pond until slurry water decanted and returned to port. Pro- vide storage for 30 days supply, reclaim as required. Coal Storage Coal received from unit train at dumping station, conveyed to coal pile for 30 days storage, reclaimed as required. Provide for dust con- trol during coal movenient. Line, Limestone Storage Materials received from train at duniping station, conveyed to storage piles (covered storage for lime), reclaimed as required. Provide for dust control during solid movement. Process Materials and Supplies Storage Hold 30 days supply of all process materials in apnropriate type of containment: pressure vessels, tankage, covered bulk storage. Assumed Depends on nodule; delivery schedule. Assumed Assumed Good 5ood Standard practice. Standard practice. Assumed 5-74 Good Standard practice. Plant Sactior.: Function.: Descr ipCion : Services - 'Cater Services Assorted Equip~".ent Provide Treated 'take-Up V.'ater. Treat nake-up vrater for cooling. potable v;ater supply, BFW and process vater surge. Criteria Source Confidence Reniarks Water Services Make-up water conditioning Lime softening to remove hardness Coinmercial from 250 ppm to 50 ppm Treatment chemical to cooling tower Commercial to clean slimes and algae etc. Treatment chemical to potable water Commercial CI2 as required. BFW treatment. Ion exchange for hard|-Commercial ness removal into NaCl regeneration Blowdovm from cooling towers: 10% to Assumed make up. Drift: 5% of make up. Good Good Good Good Fair Depends on water quality and tovei design. 5-75 Plant Section Equip-er. t Iter. r unction Descr iotion Services (con- -nu=d; Conbustio n for Dryin ~ and 'Lai: 1 3o: 1 2 r/C-as Xr satr.ent. Provide d rying gases and ~ain boil er feed to produce st ea:- and ' Burn gase s and send to dryers and bu rn ga ;es to provide e- = rgy for stear prod jction and produce ' in -house" electrical eraii Criteria Source Confidence Remarks Conbustion: (associated with each Conventional practice Good Firebox and access- operation requiring hot gas). in coal combustion. ories at each location, gas cleaning done centrally. Main boiler required to generate 810 MPPH of 185°C (11 atm) steam Calculated from pro- cess material and Good Boiler and access- ories would be of and 30,000 KW power generation from energy balance. standard design. back pressure turbines. Gas Treatment By limestone scrubbing to take 1% sulfur and other constituents in the coal out of the gas. Water sufficient to form slurry to make sludge pumpable. Off gas disposal: high velocity high stacking. Practice is conven- tional Power generation. Good Conventional practice. Fair Offgas composition a function of trace and minor ele- ment composition of coal. Exact requirements depends on local requirer.ents. 5-76 6.0 DETAILED DESCRIPTION CUPRION/AMMONIACAL LEACH PROCESS 6.1 PROCESS DESCRIPTION 6.1.1 Summary Process Description Copper, nickel and cobalt can be recovered from nodules by the Cuprion Process employing a reducing ammoniacal leach. A simplified block diagram of the process is shown in Figure 6.1. The first step in this process is a low temperature (50°C) hydrometallur- gical reduction of manganese dioxide, the major mineral in nodules, by an aqueous ammoniacal solution containing cuprous ions. The effect of this reduc- tion is to disrupt the mineral structure and release the contained metals. The metals are removed from the reduced nodules by dissolution into a strong aqueous solution of ammonia and carbon dioxide at low temperature and pressure. The metal-bearing solution decanted from the nodules is subjected to a series of purification steps in which it is contacted with an organic mediiim, which selectively and separately removes the copper and nickel from the organic fluid, and transferred to acid aqueous solutions which accumulate copper sulfate and nickel sulfate. The metal products, cathode copper and nickel, are produced from these solutions by electrodeposition. Cobalt is then recovered from the aqueous ammonia/carbon dioxide solution by contacting it with hydrogen sulfide, which precipitates the insoluble sulfides of cobalt as well as small amounts of copper, nickel, zinc and other metals not removed in previous steps. The solids are removed from the aqueous ammonia/ carbon dioxide solution and contacted with air and hot (100°C) sulfuric acid to selectively redissolve the cobalt and the small amount of nickel present. The undissolved sulfides are sold as minor products, and the cobalt and nickel are recovered from solution in powder form by selective reduction with hydrogen at high pressure (34 atm. pressure) and temperature (185°C) . The reduced nodules residue, from which the major portion (98%) of the soluble metals has been removed, is contacted with steam (at 120°C, 2 atm. pressure) to remove residual ammonia and carbon dioxide. The aqueous ammonia/car- bon dioxide mixture, from which cobalt had been removed, is also steam stripped to recover a high strength ammonia solution for recycle to the reduction step and to provide fresh wash solution for recycle to extract more metal values from freshly reduced nodules. The steam-stripped nodule residues are combined with smaller amounts of other process solid and liquid wastes and sent to containment. Plant services include facilities for generating the carbon monoxide gas used in nodule reduction, raising the necessary steam and power required for process use, supplying the make-up water and cooling required, and providing for materials handling for process materials and supplies. The hydrometallurgical reduction of nodules in an ammoniacal/ammonium car- bonate solution has been disclosed in the patent literature. While this approach differs from the pyrometallurgical reductions used in the well-known Caron 6-1 I i r <:::: 1 [^ te "^ 1/7 llCrrrr ^ 1^ i --0 M |J-1 * ■-■ - : 1. s > cSbz^ 1 1 tt? 1 II II II II II z 2 c C = " >0 5 0--:q a. £ 12 £ Zi tr^(- •^ K - — 1^ 2* iZ 3 I C=R 2g o CO u o cr 0-, u I Q O O vJ z -1^ w Pi B M J 4 lz_lz 1 I 3^^S ; I t; ?i. » ^'E < o ^ 8 < 7> II I .1 6-2 process, the basic outline for the Cuprion Process is similar. The metal separation and purif icati^^n scheme, however, is specific to nodules and is complicated by the chemical similarity of copper, nickel and cobalt. Separa- tion and purification of copper and nickel by selective extraction with organic compounds (liquid ion exchange reagents) is currently practiced in the extractive metallurgy of copper and nickel. However, in these cases the aqueous solutions contain primarily one metal, the others being treated as impurities, not products. Cobalt recovery from precipitated mixtures of nickel and cobalt sulfides derived from laterites is also currently practiced. The details of the procedures used to purify the leach solutions prior to reduction, however, would differ somewhat from those used for nodules because of the differences in amount and content of impurities. The requirement that the reduction and wash steps be carried out with solutions of differing ammonia and ammonium carbonate compositions, however, requires an additional step, raffinate stripping, not used in the pyrometal- lurgical reduction/ammoniacal leach process. The generation of carbon monoxide gas from coal or oil for the reduction of cupric ion and all other plant services represent the utilization of known technology essentially without adaptation. The basic process configuration proposed, then, is a modified version of the well established Caron process for recovering nickel from laterites. The major difference is the reduction and metals separation and purification steps, and represents an extension of elements of existing technology to a new appli- cation in nodules processing. The detailed design bases used in developing the process description are summarized in the criteria sheets in Section 6.7. A major alternative to the proposed process configuration involves the high temperature reduction of the nodules using carbon monoxide rich producer gas. This option is investigated in detail in another section of the report. Other variations in the process are possible, particularly with respect to the specific metals separation and purification schemes used. However, they would probably have only minor impact on overall inputs and outputs. Alternative approaches to the recovery of ammonia /ammonium carbonate solutions of differing strengths for recycle within the process could, however, have major impacts on the plant energy balance. Savings up to one-third in steam generation (25% in coal consumption) might be possible at the expense of purchasing up to 10 Kw of power from outside sources. In addition, it has been assumed that a major by-product of carbon monoxide production, 23 X 10 std. cu. ft. /day of hydrogen, would be burned. If markets were available, it could be" sold at the cost of making up an additional 330 tons/day coal. 6.1.2 Detailed Process Description The Cuprion Process is a three-metal process in which copper, nickel and cobalt are liberated in an ammoniacal/ammonium carbonate leach following a re- duction-leach step. Carbon monoxide is used to regenerate the cuprous ion which reduces manganese dioxide. Copper and nickel are co-extracted by liquid ion exchange reagents and selectively stripped and recovered as electrowon cathodes. Cobalt is recovered from the raffinate by precipitation with hydrogen sulfide and is recovered from the sulfide precipitate by selective leaching and hydrogen reduction along with some nickel, zinc and copper. The metal-free raffinate is steam stripped to recover a high strength ammonia solution for 6-3 recycle together with ammonia and ammonium carbonate recovered by steam stripping leach tailings, to the tailings wash step. Detailed descriptions of each segment of the process -are given in the following flowsheets. (Dwg. NC-0001-Cu, Figure 6.2) Wet nodules are reclaimed from storage and fed through a primary cage mill where they are reduced to -7/8". They then pass to a rod mill for wet grinding to -200 mesh. Oversized nodules are separated and returned via a rake classifier, hydrocyclone , water-recycle circuit. The ground nodules slurry passing through the cyclone is held in an agitated surge tank to prevent settling before being pumped to the reduction step . (Dwg. NC-0002-Cu, Figure 6.3) Thus nodule slurry is mixed with recycled strong ammonia solution and fed, with a recycled solution containing an excess of cuprous ion, to a train of agitated reduction reactors. The dilute slurry is contacted with a carbon monoxide rich gas, derived from gasified coal, the manganese dioxide is reduced and converted to manganese carbonate, and most of the value metals are solu- bilized. The heat of reaction is removed in shell and tube exchangers to main- tain the reduction temperature at 50°C, excess carbon monoxide passes to ammonia recovery, and the reduced, dilute nodule slurry is thickened with the overflow recycled and the thickened pulp passing to oxidation. (Dwg. NC-0003-Cu, Figure 6.4) The reduced nodules slurry passes to a counter current oxidizing leach in which air is sparged into the leach slurry to oxidize the cuprous ion, cobalt, and iron, precipitating the latter as an insoluble hydroxide in the manganese carbonate tailings. Liquid/solid separations are made in thickeners, which also provide residence time for leaching. Off-gases from aeration pass to ammonia recovery, leached tails to washing, and pregnant liquor to liquid ion exchange for metal separation. (Dwg. NC-0004-Cu, Figure 6.5) Metal values which have been solubilized in leaching are recovered from the tailings by washing in a counter current decantation circuit using covered thickeners. Wash liquor consists of recovered ammonia and ammonium carbonate with raffinate from cobalt recovery. Washed tailings pass to stripping for ammonia and ammonium carbonate recovery while the wash overflow passes to leaching • (Dwg. NC-0005-Cu, Figure 6.6) Pregnant liquor from leaching is filtered and passes to a three-stage, counter current liquid ion exchange extraction circuit where copper, nickel and some ammonia are removed from the pregnant liquor. Most of the ammonia is removed from the organic phase by washing with a weak aqueous ammonia solution. This ammonia is, in turn, recovered by stream stripping a portion of the scrub solution, with the vapors passing to the ammonia recovery section. Provision is made for periodically cleaning the mixer/settler units used in extraction and stripping and for recovering organic and aqueous phases for recycle. Degraded organic, dust, and other forms of "crud" are removed from the organic and incinerated. Since a small amount of cobalt is co-extracted and is not 6-4 * ^ ^ » 1 1 1 I 1^ 1 1 1 5 S I I fUUl w ^S > i W ^ e5 § w ii i ^,^ '^ ^w ^ 5^ 3) •8 \^ \ i '" ' n 5; 1 — V - ^ t ^ ^ >0 S" R iV \ I CN I V % 1 1 'T \^ 1 Pi P \^ 1 1 M is ^ r ^ \1 I \ ^t'"^ Q \\ Ci \\ 1 '^ <0 '^ ^ \ 4 ~7 '^ // 5 ^ CM ^ 5 \ fV- 1 . // -o ,5 v-x '^ lO l\ / 1 N) V y ^ «0 » >n 1 f * »- i U. ■ Afcy 1 !*1 \ 1 1 y\ -» ^ W n 5 . ^1 s. 5^^ 1 ^ ! ^ S ii; to" V) k d Vi >J Vf 6-5 6-6 ^ U) \i •J -3 ? -J ^ *^^ < i^ "i "Sa lb WP 51 o uu <0 3 ^ O Vj • Vj 7* O ^ 8 1 O :»: :2 u ^ ?*: T »<4 ^ i I ! i i 1 I I o UJ i ' I ill g M 6-8 2 I.J uu o 5^ -t— N O __ii «0 V ,1 >9 > N r; - — t t 1 (MO ^ ' tM' a , ■*! ^1 -SI "! • Ui '■ B- 1 >ni 1 O 1 1 I •?_._-♦ V ; >3 ^ t^ lln t 1 •0 ;<^ >; nS <» ! Q »n * ^ i! 1 In In > - 5 1 * in ■♦• - n M ij O 1 Q l^ fi in ;^ •u ?? : 1 C) o 5 f J ■s ■4 ui K 1 S 1 8 1 11 :3 u « o o I o 1 w M 6-9 stripped with copper and nickel, it must be removed from the liquid ion exchange reagent to prevent its buildup. This is accomplished by precipitating the cobalt, with hydrogen sulfide, from a purge stream of organic. The preci- pitated solids are washed from the organic and pass to cobalt recovery, while the purified organic is returned to the extraction loop. (Dwg. NC-0006-Cu, Figure 6.7) Tne remainder of the ammonia is removed from the loaded organic by wash- ing with a slightly acidic ammonium sulfate solution. The ammonium sulfate formed in scrubbing is purged to ammonia recovery. Nickel is then selectively stripped from the organic by counter current contact at controlled pH with depleted electrolyte from nickel electrowinnlng. Since electrowinning conven- tionally occurs at a higher temperature than operation of the liquid ion exchange circuit, the strip solution is heated/cooled passing from/ to electro- winning. Copper is then removed from the nickel-free organic by counter current stripping with depleted electrolyte from copper electrowinning. The copper electrolyte is also heated/cooled on passage from/to stripping to permit operation of electrowinning at a higher temperature. Make-up organic is added to the stripped reagent to offset degradation and soluble organic losses to pregnant liquor, and the metal-free organic is recycled to extraction. (Dwg. NC-0007-Cu, Figure 6.8) Cathode copper is recovered from the strong electrolyte from liquid ion exchange using conventional technology. Started sheets are deposited on titanium blanks from strong electrolyte in the stripper section and are removed, washed, looped, and returned to the commercial section as starters. Full-term cathodes produced in commercial section are washed, unloaded, and prepared for shipment as cathode to sale. The greater portion of the weak electrolyte is recycled to the liquid ion exchange section for stripping, but since a small amount of nickel is co-stripped with the copper and not deposited in electrowinning, it must be purged from the system. The purged electrolyte passes through purification cells where copper is removed by electrowinning to depletion. The decopperized electrolyte passes to vacuum evaporators where water is removed and nickel sulfate precipitated from the resulting highly acidic solution. The nickel sulfate is removed and sent to cobalt recovery, while the acid is returned to the process where, with make-up acid, it is used to redissolve scrap copper for return to the commercial cells for deposition. Provisions are made for recycling copper, in ammonium carbonate solution, to reduction, if required. Sufficient steam, wash water, and make-up water are added to the circuit to offset water which is vaporized and carried, with evolved oxygen, from the electrowinning cells. (Dwg. NC-0008-Cu, Figure 6.9) Nickel is recovered from the strong electrolyte by electrowinning in a manner similar to that used for copper recovery. In nickel electrowinning, however, cathode bags are used, and sodium sulfate and boric acid are added to the electrolyte to control its conductivity and pH. Dissolved organic carried from the liquid ion exchange step is removed by adsorption on activated carbon 6-10 ? *1 P 'Kj "5 V, a lu •'J W u \ liU ^•r ^ ?i i^, yj •» ^ vo 1 «< :^ "^j ^^^ ^l;; 1 o Of > ft q > ^S 1 :»: c^s- VJ :^ 4r ft. ^ 1 >| VJ ;a >■? ^ ^ »> IV) H. ...."I. \ 51 N 1 1M - i; ■N. «^ ''^ ^ '^ ^ a S 2 tvj ^r ■J ■^ ^** »- fl ^ M ■;, ■* x^ < ^ fe ^ •9 r^ o 1^ !5 ^^ to > M ^1 -- fg 5 50 - 5 '4 K M -. 5 ^ ■S 5 >^ tv4 5 «M « * \ 5\J ^ Un ^?: $'>^ !'<^ < 8; 32 a, I 2 i-'ii ■i'f 8 fi w Pi IS o _J 6-15 ^ / ;v- \ oh k > Si ^ i^ ^^ [ ^^ 1* 1? o c ^ ^ \ O ^ ^8 1 1 « ^ Vi k ^ ^ ■^ 9. i^ :^ ,^ N^ ^ ^ Vj R 0?' 6-16 The ammonia rich ammonium carbonate solution required in reduction is obtained by stripping only a portion of the ammonia contained in the raffinate stream from cobalt recovery. These ammonia rich vapors are combined with ammonia-steam vapors from the LIX ammonia recovery and lime boil steps, make-up ammonia, and ammonia rich vapors from reduction and oxidation vents and are condensed and recycled to reduction. The raffinate is then stripped further, with the ammonium carbonate recycled to tailings washing and the stripper bottoms used for process water make-up. Ammonia is recovered from the ammonium sulfate purges from cobalt recovery and the liquid ion exchange washing step by reaction with slaked lime. Steam is blown into the mixture to strip the evolved ammonia, and the vapor is condensed, with the condensate returned for recycle within the process. The gypsum slurry from the lime boil is cooled, along with the slurry from the tailings stripping, and passed to tailings surge. This is combined with process solid and liquid wastes and plant run-off, treated for pH control, as required, and pumped to the tailings impoundment area for disposal. (Dwg. NC-0011-Cu, Figure 6.12) Facilities are provided within the plant for receiving and reclaiming raw nodules, coal, lime and limestone, ammonia, and other process materials and fuel. (Dwg. NC-0012-Cu, Figure 6.13) Plant services include process and cooling water supply and treatment, steam raising and power generation, gas treatment, and reducing gas preparation. Make-up water is clarified and softened for distribution to the process, as required. Additional treatment is required for cooling tower water make-up, boiler feed water make-up, and for supplying plant potable water. The carbon monoxide required for manganese reduction is produced in a low pressure gasifier in which coal is mixed with oxygen and steam and partially combusted. Acid gases are removed from the reducing gas, following particulate removal, and are sent to the main boilers to oxidize reduced components for subsequent recovery. High strength carbon monoxide is then removed from the sweetened gases for use in reduction and the remaining gases, mainly hydrogen, are sent to the main boilers except for the amount required in cobalt reduction. These and other combustible process off-gases are burned, along with additional coal, in the main boilers to raise the required process steam and generate power required in the process. Following particulate removal, the flue gases are combined with other process off-gases and pass to gas treatment, where sulfur oxides and other constituents are removed by scrubbing with lime- stone. The scrubbed off-gases are reheated, combined with scrubbed vents from ammonia recovery, and stacked. 6-17 W o M 6-18 o uu ^^ »^ <0 ft: <0 <0 • V) 3 (0 U c o u ^ ^4 • a. >-i cd u en c o X T-t x> H 4.1 3 J O (J 4 en o en 00 CO o X! o o en en o o en -* en O O CM O en en O CM cd Cd -H 1-1 u O en < o CTv O CO o CM O . o o CM a cd •H 1-1 4.4 CJ tn < pa o CM s CM en CM -a- O aa o CM en o • Oi o CM ^»-^ (J cd •H r-l cd 4-1 o en < o U en (B CM O >> O O cd C C 1-1 S S u o o z E cd « o o o E-« -3- o « y < cd M O e (D J3 Hi B 4J O cj CD 09 o %4 a. m C o o u a •o o o T+ 4J Cd ! CO nO M 60 J3 t-i 1+-I D 3 <4-l 5j iH O CJ t/3 CO c o •T-l 4-1 >! CO u a J3 u •H en 3 U-l - cu rH M •H •iH }-l M CO d 4-1 H <-\ U^ CO e CO cu u 4J CN u-l lO CN o c iH -H fij s CO CO 4-1 la o H O in CM o CN en CN O in o in u cu 3 4-1 CO V4 cu a B cu H CO o iH <)■ in \o v^ O 0-) ro iH ^ -l OJ 4J CO CO (U o D. E o o o 3 o 0) 3 o H 6-25 CO B td (U te M CD •u (U w o o ■u M C P-i CU 3 ^ -H o 4-1 CJ u ^ o CJ w to o CU CO hJ 01 CU (U CO r^ ^ CO d CJ C 3 CO H CO o x; IZ H ffi X w 0) CU CO (X ^ CO 3 a. c 3 CO O CO O ^ w iH CO J-l O 4-1 CO !-l CD 3 w 3 Q C CO o jn w LTl 00 o rsi C3> LTl 00 CO u-i in o O m <-o r^ Pi CN LO CD r^ O- H .H vD tH (^ 3 00 C x) 3 CO 3 CO x; X CU CJ 6-26 6.4 POTENTIALLY TOXIC ELEMENT FLOWS AND DISTRIBUTION In the Cuprion/aiimioniacal leach process, the total input rate of inherently toxic elements is about 65 tons/day. Estimates of the distri- butions of the ten selected constituents were made for each process juncture by the procedure outlined in Section 4. These estimates or assumptions were then combined to produce a collection of estimated attentuation factors, as displayed in Figure 6.14. A summary listing of estimated toxic element distributions and flows to product and waste streams is presented in Table 6.6. The outputs from the Cuprion/ammoniacal leach process include gaseous emissions, liquid waste streams and solid waste streams. Liquid and solid wastes are, in some cases, combined in a single waste stream as a slurry or slime. The potentially toxic elements are distributed in varying amounts in certain of these streams. The waste streams are listed in Table 6.7 with the total flow, in tons/ day, representing the sum of the liquid, solid and toxic element components of that stream. Streams 64 and 118 will be incinerated with other combustibles, with any residue going to the tailings pond. Stream 222 will contain the material which must be removed from the off-gases to permit discharge to the atmosphere and may contain significant amounts of hazardous substances. Streams 94, 114 and 161 (comprising streams 160 and 164) will contain one or more of the potentially toxic elements. While it is possible that other process waste streams will also contain hazardous substances, such as water treatment chemicals (streams 204 and 207) , or combustion ash constituents (streams 214, 221 and 222), these have not been considered here for reasons explained in Section 4. In addition to the total flow rate and distributions of potentially toxic elements in the process streams, it is important to define their concentrations as well. This can be done by using the flow rates calculated from the attenua- tion factors summarized in Figure 6.14, and the total flow rate of the streams calculated from the process material balance. The concentrations of hazardous substances in the waste streams from the Cuprious/ ammonia leach process are summarized in Table 6.8. 6.5 PLANT SUPPORT REQUIREMENTS It is estimated that the total plant facility, including processing and support facilities, but excluding long-term waste containment areas, will require approximately 200 acres of land. Within the main processing area, the plant layout will involve equipment densities which are typical for the chemical processing industry. Separate processing buildings would be provided for the liquid ion exchange operation and for copper and nickel electrowinning and cobalt recovery. Nodule reduction, ammonia recovery, and coal gasification would be open-air installations. A boiler house/power house and banks of cool- ing towers would be located adjacent to the chemical process areas. Offices, laboratories, maintenance shops, warehouses, control houses, change house/cafe- teria, and parking area would also be located close to the main processing area. 6-27 rH ai :t3 u O u 03 ■Ul i) — ^ ;j u U_( C ^ u a O '^ 03 u q c 73 73 ^ 3 ^ ■*^, •w O y ^ _J u a u r-H —1 3 < ^ vj C d. O -a 2 J-J • r^ o 1-1 ^ 03 nj c 0) 03 * J= T3 r-l X J_) CU o- •H J-1 t3 2 ^ 4) ■H — x: CJ •H 4_) ^ CO u ^ ■i) ■Jl < r— 1 0) 4J 3 7) ■a OJ U ■r-l o — ^ ■•r - J_l OJ LJ 0) aj UJ a 71 u ^ ■-t 0) 03 OJ c 73 ^ M c 4^ 50 ^ 0) QJ ^ c > OJ « (U ~i r^ c CO cn U-l ^ cy 3 O OJ ^ o ^ ij O 73 a 73 3 u-j /^\ 0) u o o a — ( o 4-J 71 • u 73 r-^ ,— j Cu ^ o c ^ II T3 U 3 OJ S^ u r-J o 73 o •H •H JJ 03 X OJ o u 3 C3 11 u u ■a >^ iO . a Q. 71 1-1 QJ 1-1 )^ .ii 3 U^ 73 •H 73 — 1 •" CJ U - M ;=> H Pi u u ^^ H fO XI W W 3 i-J cT> in iH O O ':' o o c\ in o o cn r-i o o o ^ ■ o Tl tn jn r^ ja U < CO £1 Oi i-l O o o o o Id r-{ in a hJ > u < £h o o o o o o -' pq M CO CO c« M u O pL, PC C_) < M Q o M p^ o 6-28 I 2i o a M H t^ <: W3 H M Q M Pi PL. M PL, U SS U Q u OS tH <: PL. U M a < 1-J p^ T CM J XI CO 01 04 < <-* -H IT) Eh o> (U irt 00 in J.i OJ o rH o o o o t^ o -H >. z 4J > fl CJ ^ 0) :3> u n cc u u u O 00 .H n IT) rj- 00 03 E-< o o o o J3 r-i J3 01 en Eh CU tu < ^ r~ CT\ n in CN r~ r- in OT CO IN o o o o • • « • o >H •k 1^ CO c u u N > u CnTJ ? "S* o < U in cr> a o o \D in o o o r- to XI rH < W Eh X3 in rH (Tl CM CO IN in CM o t*- o 3 2 O :3 > u < u in -H in •^ CO in tn o o o ^ ■T tn Xi ^ < cn Eh >i 4J U in a\ rH a> >i CN O 30 > .-1 CN O O .-( ^ • « • o 10 « -H V4 2 u Q fl J > u O -o in < U rH O O o o CN O tN ^ c •H s^ •p ^ (0 04 * -u in 3 c fN u •H o O o = c tn < E-t in o o o o O O un O O > u o o in o o in in "a* o pt-: Xi Eh XI 04 in o o o o o r-l CO O O in o o > 01 -a rH CO O O in o o in CM o in o o o CD o h-t Pi P4 W Pj P^ M PL, Pi O H U C/2 CD X3 H i3 < 0] e-i 04 in i-i o^ in o o o o o < o H rH in in O O •'3' 'T o X >. u ■H Q) z > H \ 4J (H (1) (fl ai ,Q U M XI rH X3 < CO E-t 04 in H O < u (71 in in in O O r^j ^ O O O 6-30 XI en en rH X! < in E-1 On n "a* •^ 1^ (N o r— 1 O o o o O en cj rH O tn XJ rH M XI ^ < C/1 9^ X2 < W £-• XI in a\ rH CU un en ^ T o 00 CO ^ o CO CO o o o rH o o o o o o o o o o o o o o o o u u > u > u 73 D1 n 01 o in in < o tn XJ r-t XI rH m < XI O (N XI < CO Eh &4 •"a* (N O tn en ^ X3 o o ,-1 o en w ^ XI o o o w > o T! r-4 u &^ «J o U Eh (0 > o iji -a in in en T3 1— 1 in tTi tn < o Ul CO m < U CO ■v < 1— 1 tn cn CO V a o rr o o ■<3" n a la n iH T •H en o O o ^ o o o o •H tn o o (Tl P-l iH •H o O rH ro o o r^ o iH •H o o o o Q * • • • t • Q w w 4J >i tn fH &^ Sh 0) fl a o a x: C) o •u OJ u (Tl a; u Eh « ■u >1 U3 --I a; x: > h- 1 ^ u CJ 2 ^ O u M OS o 6-31 34 n C) r-l u l-l CO ^ in ■^ 03 o o O US o o (0 XI TI OT en ja U < in .-( &- n in 00 H •^ CO (N o OJ O O o •^ O o o o <-i O Q Q O is: o o 1x1 Pi M 6-32 Table 6.6 Summary Flows and Distribution of Potentially Toxic Elements Cuprion/Ammoniacal Leach Process Element Input Rate to Plant 103 lb/day Barium 76.2 Lanthanum 31.4 Vanad ium 7.6 Chromium 0.19 Silver 0.06 Cadmium 0.38 Arsenic 0.96 Antimony 0.58 Thallium Lead 3.8 9.6 Distribution as Percent Reporting to: Wastes Discharge Wastes 10^ Wastes oduct lb/day 76.2 31.4 1 7.5 1 0.18 87 0.008 80 7.6 27 0.70 9 24 0.53 2.9 7 8.9 100 100 99 99 13 20 23 91 76 93 6-33 H r T3 T3 X) T3 Xl TS T3 ■o. T3 C c c C c C C c C C C o O O o O o O O o O o o 4J •H 4-J CL, pL, •H PM Pm P-i PL, pL, p^ p^ r-i cC to 03 cd w W W CO CO en CO « >-i GO bO u 00 60 60 oo 60 60 60 Cfl 0) c C H 5-1 H !-) H X O U U - O C ^ 60 - 60 H M P-i > < > < o rH -^^^ n3 CO x-i C O O H 4J ,Q J3 C/D PM #s A CO rH <: H ^ ^ S-i TJ u U •\ #v > S-i u CO ^ #^ hJ PL, > A »• «v CO r-l CO pa H i-J o CNI o O o LO -a- en en to •H o V-i •H ^ - CO e QJ C • CU 01 C iH 3 4-1 CO O ^ ^ O •H en CO 60 4-1 CO CO J3 CO & >^ s Sj (D _o U U 4-1 l>^ u O r CO 60 S-I 14-1 •N O T3 1 3 o = CD 0) QJ QJ >s QJ XI .H !>^ Xl 60 60 60 4.1 JZ >-l 60 •H CO ^ 0) 3 T3 T3 T3 c O ^ XI 3 ;j o. V4 3 3 3 QJ CO 3 3 cr X x 3 >, U .-1 1— t iH a. QJ .H rH •H CO CO .H H : CO cyD C/2 CO hJ CO CO hJ <; <: CO CM ON CM CO CO cn Cl- pi< cn e e 60 3 3 C CO CO •H >J 60 60 60 •H (U (0 C c C CO •u 0) •r-t •H •H ,H 4-1 00 CO U 4J C c C CO CJ & QJ CO c c c > CO •H ,Q CO •H •H •H O •H C XI C C cn ^ S C & S s e > QJ QJ o O i-i 3 O O O O Q) 4-1 QJ ■iH •H 01 U u-i •H S-i U VJ C^ r. 60 iH 14-1 1+-J 4-1 4-1 .H CJ o 4-1 4J 4-1 4-1 ^ C •H O CO CO •H CO o O CJ O CJ cn •H O CO »* N a O c QJ 0) QJ QJ -H CO & pq >-i •H •H pa cn •iH CO rH -H -H C IS P. u QJ C 14-1 • CO 60 W W W CO •H QJ QJ iH O •H C 60 •H X 60 o 1-4 e 4-1 •H •H cn •H M-l Vj 1—1 3 •H •H ^ u 4J •H CO O Q) CO CO >H O J U s Z O u cn ►J :2 pa XI o s o S CO QJ i-t ' H C tH •H C ». c . ,— ^ (-1 B !-i a 3 (i iH ^-^ CO . B U cx u Ci 3 ^-^ CO o O) Cfl tH bS) C B •H cfl rH 0) ■H >-l Cfl ;-! H CO cn cn Q) a o ^1 o ■H O CNJ CM O O ^ O CTn CN VO -^ O og ^ O t^ CM r^ .H rH CT^ ^ CN \0 CN CN CN CN CN 00 so rH -J- o o O O O O cn LO CN cn cn in C3^ c>0 o o 00 CN a\ 00 00 e d e B >^ B c 3 3 S o C 3 B cfl ■H •H u 3 •H O •H 3 X T3 B 0) •H C B tH •H 4-J cfl o > B 0) •H ^ 13 S-< c C >-i rH T3 cn ■U cfl CO cfl cfl cfl -d •H cfl M C -C CD pq hJ > a CO U < <; H hJ 6-35 While the aforementioned facilities would require approximately 20 to 50 acres of land, the plant layout is dominated by the requirement to provide materials storage areas and for the thickeners in the leaching and washing section. Approximately 25 acres would be required for the thickeners, nodule storage, and decant ponds. The coal, lime, and limestone storage area and a plant run-off and emergency waste-storage area would require at least 50 acres, including margins of relatively flat terrain. No unusually tall structures would be required in the process areas, except perhaps for the main stack, which could be several hundred feet high, depending on local meteorological conditions. The materials movement in and out of the plant is significant, even exclud- ing raw nodules, waste products, and oxygen. In excess of 2,200 tons/day of fuels and process materials are brought in, and 220 tons/day of products are shipped out. Thus, the transportation network supporting plant operations must include highway and rail connections capable of supporting considerable traffic. The coal, lime, limestone, ammonia, hydrogen sulfide, and sulfuric acid would be delivered by rail, and copper and nickel product would be shipped by rail. Most other materials and supplies would be trucked, except that 660 tons/day of oxygen would probably be purchased from an adjacent or nearby facility and delivered to the plant by pipeline. Plant operations will be on a 3-shift, 365 day per year basis. It is estimated that approximately 500 people will be employed in the facility, includ- ing operating and maintenance labor, supervision, and plant general and adminis- trative personnel. The mix of skill levels would be typical of a chemical processing plant, with approximately one-third being skilled tradesmen and lead operators, one-third unskilled laborers, and one-third professional/manage- ment and clerical/support personnel. 6.6 ACCIDENTAL RELEASES The Cuprion Process is similar in design and materials input to the ammonia leaching process (see Section 5.6). However, a high-pressure gasifi- cation unit is not used. A moderately high-pressure, carbon dioxide removal unit (30 atm.) and a low-pressure hydrogen removal system (venting into the boiler furnace) are used instead. Hazardous materials in storage include 500 tons of ammonia and hydrogen sulfide and up to 15 tons of chlorine, each based upon a 45 to 60 day supply. As much as 100,000 gallons of concentrated sulfuric acid could be held in the materials supply. The liquid volume of acid solution could be increased if supply were based upon diluted acid. The risks of large pollutant releases are similar to those from the ammonia process. Approximately 25 million gallons of liquid would be held in the process inventory, and the toxic gases would be held in storage. The fire-explosion hazards would be normal to an industrial facility which utilizes flammables and fuels. The flammables include the kerosene used as a LIX reagent solvent, 100,000 tons of coal, low-quality fuel gases, and hydrogen. The fuel gases have limited opportunity for explosion outside their feed lines, but the hydrogen venting into the furnace constitutes a local hazard. This operation is performed routinely in refineries, but still requires delicate design and surveillance . 6-36 The probability of major pollutant releases are assessed as equivalent to those for an ammonia process plant : 10"^ per year for liquid releases such as the kerosene, slurry, or acid solutions of salts; 10~5 per year for major gas releases of ammonia and hydrogen sulfide. Although a major gas release could produce short-term hazardous concentrations of pollutants at distances exceeding 9 miles, the probability of exposing a given location at that distance is at least two orders of magnitude less than that for the occurrence of a release. In this context, a failure rate of 10~ per year is the same level of reliability required of the radiological system of a nuclear power plant . The level of concern for the chlorine storage release would extent out to approximately 2 miles downwind from the plant. The amount of chlorine in storage is less than that carried in a typical rail tank car or stored at a water treatment plant. 6.7 CRITERIA SHEETS 6-37 6.7 CRITERIA SHEETS Plant Section: Equipment Item: Function: Description: Ore Preparation Cage Mill, Rod Mill, Rake Classifier, Surge Tank Grind nodules for suspension in leach solution. As-received nodules are initially ground in a cage mill then passed to a rod mill for wet grinding. Slurry from the rod mill is rake classified with over size returning to rod mill while remainder of slurry goes on to processing. Criteria Source Confidence Remarks Feed Rate: 12,500 TPD of wet nodules with a surface water content of 4 wt %. Nodules composition based on Tables 4.2 to 4.4 Feed Size: 4 mesh Range: 0.5-25 cm dredged Mean: 3 cm (1-1/8") Cage Mill Reduction = 6 mesh U.S. Sieve Series Rod Mill Reduction 60 mesh U.S. Sieve Series Percent loss of iron scrap and refuse: .005 wt % of the raw feed nodules Rake classifier operated such that 60 wt % of ground nodules exit at -200 mesh. Hydraulic cyclone operated such that -200 mesh nodules pass to surge tank Slurry density in surge tank, 50% solids. Slurry temperature: 25°C Aeration rate in surge tank: at least 3.1 ft-^/(ft2 tank cross-section) (min) Various Brook and Prosser Earner et al. Earner et al. Queneau Assume Assumed Perry's 4th ed. pp. 18-71. 6-38 Fair Fair Fair Fair Fair Fair Poor Poor Poor Plant Section: (continued) Ore Preparation (continued) Criteria Source Confidence Results Circulating Pump: 150% circulation load. Cyclone Feed Pump: 100 ft head Cyclone Underflow Pump: 30 ft head Slurry Peed Pump: 50 ft head Vent Air: Saturated with NH^ , CO^ , H2O @ 25°C, 1 atm. Assumed Assumed 6-39 Poor Poor Plant Section: Reduction /Leach Equipment Item: Reactors (316 S/S), Clarifier, Make-TJp Tank Function: Reduction and conversion of Mn02 to allow value metals of nodule s to become soluble in a strong ammoniacal ammonium car bonate solu txon. Slurried nodules are multi-fed- by slurry pumping to f our of six stainless steel agitator reactors. CO ga s is sparged from the b ottom to promote chemical reaction. Unsoluble material is separated b y a clarifier and sent to "washing. Some liquor is recycled with mak e- up NH3/CO2. Description: Criteria Source Confidence Remarks Six (316 stainless steel) agitator reactors Residence time per reactor: 20 min Cascade design Reactor temperature: 50°C Pressure: 50 psi (100-5) 100 g/l NH3 25 g/Jl at steady state CO: @ 50 psi (first stage) % Composition: 93.6 CO 4. A H2 2.0 H2O 96% stoichiometric CO utilized Ammonia losses less than 5% % Reduction: ■ Manganese - 96 (at minimum) Iron - none % Extraction: (minimum) Ni - 90 Cu - 80 (in reduction) : 90 (overall) Co - 70 Mo - 80 Zn - 40 Earner et al. Earner et al. Earner et al. Earner et al. Earner et al. Earner et al. Earner et al. Calculated Earner et al. Calculated Earner et al. Szabo Redman Redman 6-40 Fair Fair Fair Fair Fair Fair Fair Good Fair Fair Good Poor Lower Cu extraction will require make up of BMC to maintain cuprous ion concentra- tion. Plant Section: (continued) Reduction/Leach Criteria Source Confidence Results pH maintained alkaline at 10.8 Slurry Density Leaving Clarifier: 40% solids (dry basis) Earner et al. Earner et al. Good 6-41 Plant Section: Equipment Item: Function: Description: Oxidation-Leaching Oxidation Tank, Thickner Settlers Oxidation of cuprous ion to cupric ion and extraction value meta ls. Clarifer underflow combined with second stage leach liquor is oxigenated with air to convert cuprous ion to cupric ion. The The slurry is then further leach-washed to remove soluble value metals. Criteria Source Confidence Remarks Separated leached pulp in covered thickness of conventional design. 3 3 Aeration Rate: 20 ft /ft slurry Thickener Settler Residence Time: 12 hr/each settler Pregnant Liquor Composition: a) 65 g/J, CO2 b) 4-8 g/S. Cu c) 5-10 g/X, Mi Standard Practice Assumed Earner et al. Good Poor Fair Earner et al. Fair 6-42 Plant Section: Wash/Leaching Equipryient Item: Thickener-Settlers Function: Recovery of soluble metals from insoluble pulp. Description: Insoluble pulp is counter-currently washed with a leaching solution to remove soluble entrained metals. Criteria Source Confidence Remarks Wash leached pulp in covered thickness of conventional design. Underflow Density: 50% solids Unit area requirement: 1.0 m /ton/day Flocculant (per stage): 0.01 kg/kg (3 1% solution. Wash rates: 2.0 kg/kg of liquor in underflow. Wash Recovery: 98% of entering solubilized metals. Wash Liquor Composition: 100 g/i NH3 100 gl I CO2 Wash Temperature: 40°C Pressure: 1 atm Pregnant Liquor Surge: 2 hr holding time Earner et al. Assumed Assumed Assumed Design Variable assumed. Calculated Earner et al. Assumed Allowance 6-43 Good Fair Poor Poor Poor Fair Good Fair Poor Also reported for laterite ore wash at Nicaro , Cuba. Similar materials will wash at this density. Based on formula from Perry's Hand- book, 3rd ed. Plant Section: Equipnent Item: Function : Description: Liquid Ion Exchange/Extraction Extractions Extract copper and nickel from airanoniacal leach liquors using organic fluid ion exchange reagents. Criteria Source Confidence Remarks Polish filtration Sand filters for solids removal 10 ppm Surge holding time : Pregnant liquor and raffinate 1 hr each. Extractions No. of stages required: 3 External O/A: 1 Internal O/A: 1 Mixing time: 3 min 2 Settling rate: .2 £/min/m % Extraction Ni 99.9 Cu 99.9 NH. Co-^ 5 1 Zn 10 Ammonia Wash - Primary No. of stages required: 2 Internal O/A; 1 External O/A: 3.3 Mixing time: 3.75 min Mixing power : Settling rate: .15 £/min/r Standard practice Poor Allowance Tests on nodules leach liquor Standard Agarwal-Beecher etc. Good Fair Test on nodule leach liquor ?air Jood 6-44 Solids inlet loading, particle size, maxi- mum allowable solid contents unknown. ■lay be specific to solution tested, short :erm test. Requires ]o fully oxidized. i Plant Section: Liquid Ion Exchange - Extractions (continued ) (continued) Criteria Source Confidence Results Residual NH3 in organic 0.1 g/Z NKo recovery from scrub solution by steam stripping 30% of scrub solution. Net removal of other constituents: . . . None Standard practice Crud Removal Crud removed from mixer/settler in- terface at 50:50 aqueous: organic. Carrying 1 w/% solids and degraded organic. 100% removal of solids/ organic by polish f iltration/cen- trif ugations . Co Stripping (precipitation) H2S required: 200% of stoichio- metric requirements for precipi- tations of Cu, Ni, Co, Zn. Precipitation Conditions : Temperature: 40''C pressure: 1 atm Residence time: 4 hr Vash/neutralization : 0/A = 1 with 100 g/i NH3 solution Solvent Recovery Recover entrained organic from raffi- nates by air floatation in standard cells. Residual organic: 50 ppm Promoter requirements: none Residence time: 20 min Reported operating practice on copper recovery systems Lab tests on organic Bxtractory simulated Leach liquors Assumed Good Poor Poor Fair Reported operating practice on copper recovery systen Ass'jmed 6-45 Poor Fair Poor Recover ea NH3 at 30% Assunes build up to steady state level in closed systems Rate and effect of crua build-up and detailed compositions unknown. Effectiveness over long term and det-'^J. of process requii ments not known. Solubility losses inknown. Assuaed solubility liTLit Plane Section Liquid Ion Exchange Operation: Stripping Function: Description: Strip nickel and copper froni organic ion exchange reagent Criteria Source Confidence Ren.arks Ammonia Scrub - Secondary No. of stages required: 2 Internal 0/A: 1 External 0/A: 1 Mixing time: 3 min Settling rate: .2 2./min/m2 Residual NH3 in organic .01 g/2. Scrub solution compositions 200 %ll (NH3)2S04 1 g/i H2SO4 Net removal of other constituents Mi - negligible Cu - negligible Zn - negligible Ni Stripping/Cu Stripping No. of stages required: 3/2 Internal 0/A: 1/1 External 0/A: 5/3.1 Mixing time, min: 9/3.75 Settling rate: % stripping for Ni 98. 8/. 001 Cu < .004/87 Co 0.3/0.2 Zn -/lOO .15 5./inin/m' Test on nodule leach liquor Good Assumed Allot-anc; Test on nodule leach liquor Fair Poor Good pH control will be required to prevent coextraction of Ni. 6-46 Plant Section: Liquid Ion Exchange - Stripping (continued) Criteria Confidence Results Weak Electrolyte H2SO4 g/Jl 40/160 Ni %ll 50/10 Cu g/S, < .001/40 Others /Zn, 5 Temp. °C 40 Surge Holding Time Electrolytes: 1 hr Organic: 4 hr Solvent Recovery, Crud Removal Same as extractions. Allowance 6-47 Operation : Copper Electrowinning Cor-.oor'.enr: : Starter Sheet Section Criteria Scares I Confidence 1. Blank material - titaniuia, 1/8" 2. Sheet size (area) - 1.00 n 3. Sheet weight - 5.0 kg 4. Current density - 200 kfwT- 5. Current efficiency t 92% 6. 7. 8. 9. 10. 11. 12. 13. 14. 15. 16. Deposition time — 23 h Cathode spacing, C/C - 0.11 m Cell voltage - 2.2 V Power consuEption - 2.0 Mv'h/kg Loop requirements — 1 sheet/5 Electrolyte circ, rate - 40 X./h/m' Electrolyte teaperature - 45°C D enlisting provisions - plastic sheet Anode material - Pb/8% Sb Cu ±a/Cu^\ut - 55/53 k/1 H^SO, in/H,SO. out - 160/163 g/1 2 4 2 4 — Industry trar.d Industry typical Industry typical Design feature Reasonable value Industry typical Industry typical Reasonable value Industry typical Industry typical Reasonable value Reasonable value Cyrus-Bagdad Industry range Design feature LIX circuit 6-48 saxr Good Fair Fair Fair Good Good Good Good Good Fair Fair Fair Fair Fair Fair '-Without loops Within range of practice. Variations in. practice. 1-day cycle. Variations in practice. D.C. electric only. Exclusive of scrapped sheets. Exclusive of flow- through. Variations in practice. Foaaing agents not used with LIX. Improved anodes under development. Within range of practice. iCH_SO, )determined b" 2 4 .(Cu^-^) Operation : Copper Electrowinning Consonant: Cori'inercial Section Criteria 1. Cathode size (area) - 1.00 in 2. Cathode weight (total) 68 kg 2 3. Current density - 180 A/m 4. 6. 7. 8. 9. 10. 11. 12. 13. 14. Current efficiency - 93% Cathode spacing, C/C - 0.11 ra Cell voltage - 2.1 V Power consumption - 2.0 kU'h/kg Electrolyte circ. rate - 50 l/h/m Electrolyte temperature - 50°C Delisting provisions - poly - propylene balls Anode material - Pb/8% Sb Cu""in/Cu^"*'out - 53/4U g/1 K.SO, in/H„S0.out - lb3/175 g/1 ^2=^4 ' 2'^"4^ Confidence Industry typical Detected by cathode cycle Design feature Design feature Industry typical Reasonable value Industry typical Reasonable value Reasonable value Most coHzon practice Industry range Design feature LIX circuit 6-49 Fair Fair Fair Fair Good Good Good Fair Fair Good Fair Fair Fair includes v;t..of starter sheet Vithin. range of practice. 7-day cycle Variations ia practice D.C. electric only Exclusive of flow- through Allowance for heat- I-proved anodes u^der development Vithin range of practice i(H2S0,)determined b by A(Cu2+) Opera:: ion : Kickel Eleccrowinning ^or.Dor'.ent: : Starter Sheet Section Criteria | Source '' Confidence 1 Re -arks 1. Blank material - stainless steel Good 2. Sheet size (area) - 0.86 m^ Fair 3. Sheet weight - 7.5 kg Fair Vithout loops 4. Current density - 180 A/m^ Design feature Fair 5. Current efficiency - 96% Fair Sensitive to operat ing conditions. 6. Deposition time - 46 h Fair 2-day cycle. 7. Cathode spacing, C/C - 0.165 m Fair 8. Cell voltage - 3.4 V Fair Determined by O.D. spacing. 9. Power consunption - 3.75 kWh/kg Derived quantity Fair D.C. electrical only 10. Loop requirenients - 1 sheet/5 Fair ZKclusive of scrapped sheets M. Electrolyte flowthrough - 31 1/h Fair Per cathode bug. 12. Electrolyte teniperature - 60°C Fair 13. Denistins provisions - anode hoods Fair Fair 14. Anode material - rolled lead Chemical lead cefonns 15. Ki^"'"in/Ni~'^out - 75/65 g/1 Design feature Fair 16. H^sn^in - pH 3.5 Design feature Good Approx. optimum pH 17. V sn out - 17 g/1 2 4 Process dependent Fair Determined byA(N"i2-) .18. Ka^SO, concentration - 100 g/1 Design feature Fair Required for con- ductivity. 19. K^BO^ concentration - 15 g/1 Design feature Fair P.eauired for pH con- trol in catholyte. *E:' .cept as noted, Outokurapu Oy P' •actice as of 1954. 6-50 Operation : Nickel Electrowinning Conpor.eaS : Conmeircial Section Criteria 1 Source Confidence P.erarks 2 1. Cathode size (area) - 0.86 m Fair 2. Cathode weight (total) - 61 kg Fair Includes vt of starter sheet 2 3. Current density - 180 A/m Design feature Fair 4. Current efficiency - 93% Judgement Fair Sensitive to operat- ing conditions. 5. Deoosition tiiae - 164 h Fair 7-day cycle. 6. Cathode spacing, C/C - 0.165 m Fair 7. Cell voltage - 3.4 V Fair Determined by O.D. spacing 8. Power consunption - 3.86 k^'Jh/kg Derived quantity Fair D.C. electrical only 9. Electrolyte flowthrough - 11 1/h Fair Per cathode bug 10. Electrolyte temperature - 62°C Fair 11. Denisting provisions - anode Fair hoods 12, Anode ciaterial - rolled lead Fair Rolled lead deforms 13. Ni^^in/Nl out - 75/46 g/1 Design feature Fair Gives net A(Ni ) = 25 g/1. 14. H^sn^in - pH 3.5 15. -H^Sn^out - 52 g/1 Design feature Process dependent Good Fair Acprox. optimum pH. .2+ Determined byi(Ni ) 16. NaTSO, concentration - 100 g/1 Design feature Fair Recuired for con- ductivity 17. H^BO- concentration - 15 g/1 Design feature Fair Required for pH con- trol in catholyte. *E: f V r .^-- S,:..'] "-.'..'-- 2=" .•.-■-;£ 5 Provi'i-i- iiurr^s. C^:^^--:-:/ f-.r ?; = '.-: O^-'^rl-j ^eouir^j-iris ?roce-s - s ' e ^ ial". a'lcl ^lor-Ii-:^ -.'re recei/ed fcon various source?. , ri--l'J in vio r ?:-',=■ , '-'^r.-'. r ^^-z .'"1 ='/?! a^ rsquirsd. Criteria Confidence RenarJ-.s ^Joduies storage: Modules rsc^iv^d as 30% slurry and hald vat in cor.- tairoeat pond until slurry vatsr decanted and returned to pr.rt. ?r'. vide storage for 30 days sj;:-ply, reclain as required into surfaca ZLoisture. Coal Storage Coal received fron unit train at dumping station, conveyed tc coal pile for 30 days storage, reclai-'.c^d as required. Provide for dust con- trol during coal :noveT:;iat. Line, Lioestona Storage Materials received fro:a train at duf.oing station, conveyed to storage piles- (covered storage for lir.e) , reclaimed as required. Pro-/ida for dust control during solid covenant. Process Materials and Supplies Storage ] Hold 30 days supply of all process materials in appropriate type of containi-ent: pressure vessels, tankage, covered bulk, storage. iced As surged Ass Limed Good Dapends or. r.cduiS5 delivery scnscuj-E. Standard practice. Standard practice Standard practice. 6-58 Plant Section Function Description Services - Uater Services Assorted t-auionent Provide Treated Make-Up '." Treat make-uo v.'ater for cooling, Dcrable vater supply, d': process water surge. Criteria Source Confidence Pv.er-.arl-:: Water Services >'ake-up water conditioning Lime softening to remove hardness from 250 ppm to 50 ppm Treatnient chemical to cooling tower to clean slines and algae etc. Treattr.ent chemical to potable water CI2 as required. Cotnniercial Commercial Commercial BFW treatment. Ion exchange for hardj-Commercial ness removal into NaCl regeneration. Blowdovm from cooling towers: 10 to nake up . Drift: 5% of make up. Assumed Good Good Good Good Depends on water quality and tove: design. 6-59 Plant Sec-icr Lcui~-sr.c leer Fur.ct ior. : Descriotioa: Services (ccr.-ir.ued) Cor^.bustion for Dryins ari M^ir. 1: .:-/:as Treatr^enz. Provide drying gases av.d - = i:^, bzi .- fe- -0 produce st-- ar.d : Burn gases and send to crysrs ar.d D-n i-^ses to prcvice -argy for stear'. production and produce 'in-hc'jse" electricii gs-era-i Criteria Source 3nf idance Re-ar-.i Coc:bustion: (associated with each operation requiring hot gas). Main boiler reauired to generate 1180 X 103 Ibs/'hr of 300° C (82 atm) steam and 51,500 Kw power generation from back pressure turbines. Ga; rreatment By liTiestone scrubbing to take 1% sulfur and other constituents in the coal out of the gas. V.'atar sufficient to form slurry to nake sludge punpable. Off gas disposal: high velocity high stacking. Conventional practic in coal cor^busticn. Calculated fron pro- cess material and energy balance. Practice is conven- tional Pcver generation. Good Good Conventional practice. Good Firebox ar.i acres; at each location., gas cleaning dor.e centrally. Boiler and accessor!; vould be of standard design. Cffgas composition a function of trace and minor ele- nent cortiposition of coal. i;-:act requirezier.t: depends on local requirer.ents. 6-60 7 DETAILED DESCRIPTION HIGH TEMPERATURE SULFURIC ACID LEACH PROCESS 7.1 PROCESS DESCRIPTION 7.1.1 Summary Process Description Nickel, copper and cobalt can be recovered from nodules by a "pressure cooking" process at high temperatures and pressure in an aqueous sulfuric acid solution. A simplified block diagram of this process is shown in Figure 7.1. The first step in this process is a high temperature (470°F) and pressure (35 atm.) "cooking" of the ground nodules. Nearly the entire quantity of major minerals of value in the nodules become dissolved in the hot, strong (30%) sulfuric acid solution. After cooling, the nodule residue and solution are separated by decantation. Water that is used to wash the residue free of acid and metals is combined with the solution, and the residue is sent to a containment area. The metal-bearing solution, which contains residual acid, passes to a pH-adjustment step to make the solution more suitable for further processing. The solution is then contacted with an organic medium to selectively remove the copper and nickel. The extracted nickel and copper are separately and selectively removed from their respective organic fluids and transferred to acid aqueous solutions, which accumulate nickel and copper sulfate, respective- ly. The metal products, cathode nickel and copper, are produced from these solutions by electrodeposition. The cobalt constituent of the solution is then recovered by contacting with hydrogen sulfide, which causes insoluble sulfides of cobalt as well as small amounts of copper, nickel, zinc and other metals not removed in previous steps to precipitate from solution. The solid residue is removed from solution and contacted with air and hot (100°C) sulfuric acid to selectively redissolve the cobalt and the small amount of nickel present. The undissolved sulfides are sold as minor products, and the cobalt and nickel are recovered from solution in powder form by selective metal reduction through use of hydrogen gas at high pressure (34 atm. pressure) and temperature (185°C) . The solution, depleted of copper, nickel and cobalt, is chemically treated to recover the ammonia introduced during the neutralization step. The recovered ammonia is recycled for use in the process, and the ammonia-free solution is returned to wash freshly leached nodules. Plant services include facilities for generating the necessary steam and part of the power required for process use, supplying the make-up water and cooling required, and providing for materials handling for process materials and supplies. A major alternative to the proposed process configuration involves a low temperature, low pressure treatment of slurried nodules in acid solution. There are, however, major drawbacks, such as solubilization of an undesirable 7-1 7-2 amount of iron that would need to be removed, longer treatment time, and lower recovery of nickel, copper and cobalt. A second alternative would involve the evaporation of metal-free ammonia solutions and crystallization of fertilizer-grade ammonium sulfate for sale, rather than recovery of ammonia to the process with disposal of gypsum. This alternative, while technically feasible, would need to be studied carefully in view of its overall economics in the metals production plant and for its impact on the local fertilizer market. Other variations in the process are possible, particularly with respect to the specific metals separation and purification schemes used. How- ever, they would likely have only minor impact on overall process inputs and outputs. Descriptions of high temperature and pressure acid treatment of nodules has had limited exposure in the literature, with most descriptions occurring in foreign patents and papers. A basically similar process for treatment of nickel laterite ores, however, has been used at Moa Bay, Cuba. This config- uration proposed is an update using currently available technology of Moa Bay operations. The metal separation and purification scheme, however, is specific to nodules and is complicated by the chemical similarity of copper, nickel and cobalt. Separation and purification of copper and nickel by selective extrac- tion with organic compounds (liquid ion exchange reagents) is currently practiced in the extractive metallurgy of nickel and copper. However, in these cases the aqueous solutions contain primarily one metal, the others being treated as Impurities, not products. Cobalt recovery from precipitated mixtures of nickel and cobalt sulfides derived from laterites is also currently practiced. The details of the procedures used to purify the leach solutions prior to cobalt reduction, however, would differ somewhat from those used for nodules because of the differences in amount and content of impurities. The generation of steam and all other plant services represent the utilization of known technology essentially without adaptation. The basic process configuration proposed, then, is a modification of the Moa Bay process for recovering nickel from laterites. The major difference is in the metals separation and purification steps and represents an extension of methods of existing technology to a new application in nodules processing. The detailed design bases used in developing the process description are summarized in the criteria sheets in Section 7.7. 7.1.2 Detailed Process Description The high temperature sulfuric acid leach process is a three-metal process in which copper, nickel and cobalt are selectively leached from the nodules by strong sulfuric acid at high temperature. After separation of the leaching residue and metalliferous solution by washing, the copper and nickel are co-extracted by liquid ion exchange reagents and selectively stripped and recovered as electrolytic cathodes. Cobalt is recovered from the raffinate by precipitation with hydrogen sulfide and is recovered from the sulfide preci- pitate by selective leaching and hydrogen reduction, along with some nickel, copper and zinc. The metal-free raffinate liquor, basically water, is recycled to the washing process. Ammonia consumed in the process is recovered and re- cycled to the process for use in pH control. Detail descriptions of each segment of the process are given in the following flow sheets. 7-3 (Dwg. NC-0001-SU, Figure 7.2) Wet nodules are reclaimed from storage and fed through a primary cage mill where they are reduced to -7/8". They then pass to a water-filled rod mill for a second and final size reduction to a powdery -325 mesh. Oversized nodules that escape the initial rod milling are separated and returned via a rake classifier, hydrocyclone, water-recycle circuit. The ground nodules slurry passing through the cyclone is held in an air-agitated surge tank to prevent settling until it can be fed to the high temperature sulfuric acid leach reactors. (Dwg. NC-0002-SU, Figure 7.3) Slurried nodules enter a steam-sparged preheater along with pregnant liquor, which is recycled leach solution containing the metal values. The solution at 105°C passes through a second steam-operated heat exchange step, reaching 170°C by use of sparged steam. A final temperature of 245°C is reached within the leach reactors, where contact with sulfuric acid is made. After a set residence time within the reactors the slurry passes through two pressure and one vacuum flash stage for steam heat recovery and slurry tempera- ture reduction to 50°C. (Dwg. NC-0003-SU, Figure 7.4) Metal values which have been solubilized in acid leaching are recovered from the tailings by washing in a counter current decantation circuit. Adequate residence time in this circuit allows recycled water to wash the solubilized metal values from the precipitated and unreacted solids which pass from the system to a waste disposal area. The solubilized metal values solution (pregnant liquor) proceeds to an initial pH adjusting step in pre- paration for selective metal removal. (Dwg. NC-0004-SU, Figure 7.5) The pH adjustment process removes the necessary residual acid concentra- tion of the pregnant liquor through the use of calcium carbonate, i.e., lime- stone. Calcium sulfate, the precipitated product of neutralization, is directed to the wash circuit, while the neutralized pregnant liquor overflow of the clarifier flows to the copper liquid ion exchange process. (Dwg. NC-0005-SU, Figure 7.6) The filtered pregnant liquor passes to a three-stage counter current aqueous-organic liquid ion exchange circuit with interstage ammonia pH adjust- ment. Nearly complete removal in an acidic medium of the copper metal from the aqueous to the organic exchange liquid is accomplished along with trace exchanges of the nickel, cobalt, and zinc metals. The separated and loaded organic stream is counter currently stripped of its copper value with depleted electrolyte from copper electrowinning . The copper electrolyte is heated/ cooled on passage from/to stripping to permit operation of electrowinning at a higher temperature. Make-up organic is added to the electrolyte-stripped reagent to offset degradation and soluble organic losses to pregnant liquor. 7-4 y m (M Ok Oh m vD < o vO :3 tf> 1 1 Of ^ U 2 7-5 I 1 fUUl 1^ N 03 CD M (a fM ^ ft/ Vj <0 ^ J^ 1 »y «VJ 1 , 1^ ft: 1 1 iru <^ H, ^ ^ <3; ■ , !"- ^ ^ \^ /^ - / [ 0) ^^ ^€) ■J T\ 1 y^-y" ^ t^ '^ y a^|^ to 5! , , CT) 1 N < Vi "O ■5^ ! ^ . «o 1 fM - § /i^ <^^l 11 - vn X ^ A^ _^ \ ^^r ■^ ^ I J2j^ - ' ^ if >y 5 /fK \ ^-< ' f . no ^ S o -^^ ' ^ ^ - '*• '^^i \ K ^ k ! ^ § •^ *> ^ (*1 Ni 1 & §> f<1 o » '^ S "s -— <0^ ^ «> i ^ Si k R. p ro >J> ( vj ^ rJ §5 ^ ^ 5 1 u \ "■ N4 >') k ti k 1 ^ si v> 7-7 7-9 A continuous purge to and recycle from a trace metal stripping step is operated to prevent a build-up. The aqueous raffinate stream containing the nickel, cobalt and trace metal values is sent to a neutralizing step prior to nickel recovery. (Dwg. NC-0006-SU, Figure 7.7) Cathode copper is recovered from the liquid ion exchange strong electro- lyte using conventional technology. In the stripper section, copper is deposited on titanium starter blanks, which are then removed, washed, looped and returned to the commercial section as starters. Full-time cathodes pro- duced in the commercial section are washed, unloaded and prepared for shipment as cathode to sale. The major portion of the weak electrolyte is recycled to the liquid ion exchange section for stripping, while a small purge stream is recycled to the pH adjustment process to prevent a build-up of undesirable co-stripped metals other than copper. Make-up acid, used to redissolve scrap copper for return to the commercial cells for deposition, replenishes the weak electrolyte (a strong acid solution) for return to copper stripping. Suffi- cient steam, wash water, and make-up water are added to the circuit to offset water which vaporizes and is carried, with evolved oxygen, from the electro- winning cells. (Dwg. NC-0007-SU, Figure 7.8) The nickel-bearing raffinate from the copper ion exchange process contains an excess of acid, which must be neutralized with ammonia and made slightly basic before the nickel value can be removed. In addition to neutralization, tank aeration causes precipitation of iron, manganese, magnesium, and aluminum initially extracted from the nodules and carried with the raffinate. The precipitated metals are settled, centrifuge removed, and returned to the wash circuit before leaving the process as waste. The overflow raffinate containing some entrained precipitate overflows the settler and passes on to nickel liquid ion exchange. (Dwg. NC-0008-SU, Figure 7.9) The neutralized raffinate is filtered of entrained precipitate, then stripped of nickel by contact in three stages of pH-controlled counter current stripping with organic ion exchange liquid. Nickel and other trace metal and ammonia values are exchanged. The nickel-stripped aqueous phase, containing primarily cobalt, is directed to cobalt processing. The organic phase must be refined of entrained aqueous, dissolved ammonia, and extracted nickel and trace metals. The entrained aqueous is removed in a liquid separation step. The ammonia requires removal by two stages of recovery. In the initial stage, the organic is counter currently washed in two stages, then this aqueous phase is steam stripped of some ammonia. 7-10 7-11 o m \n <^ fe '^ ^ " ^ I I to 2 o M 7-12 ► 7-13 (Dwg. NC-0009-SU, Figure 7.10) The partially ammonia-stripped organic passes to a second step for reaction with sulfuric acid to draw the remaining ammonia into an aqueous phase. The organic aqueous phases are counter currently contacted and separated in two stages. The aqueous phase returns to nickel extraction. The organic phase, containing nickel and trace metal impurities, enters two counter current stages of nickel stripping with weak electrolyte from the final nickel-producing cathode stage. This electrolyte is heated to and cooled from nickel cathode or electrowinning operations which must be slightly warmer. The organic, having released nickel and some trace metals to the electrolyte, is replenished of losses with fresh organic. A small amount of organic is purged to a trace metals removal step to prevent a build-up. The "cleaned" organic purge and make-up organic are combined with the nickel-free organic for recycle to the nickel ion exchange step. (Dwg. NC-0010-SU, Figure 7.11) The strong electrolyte is chemically conditioned with sodium sulfate and boric acid to control its conductivity and pH. Dissolved organic, carried from the liquid ion exchange step, is absorbed on activated carbon prior to any electrolyte passing to electrowinning. Nickel is recovered in a manner similar to that used for copper recovery. The starter sheets are pickled in sulfuric acid prior to use in the commercial cells, and nickel scrap is redissolved in ammonia-containing raffinate and recycled to the ion exchange process. The electrolyte purge required to remove impurities from the electrowinning circuit passes to raffinate neutralization. The weak electrolyte, strongly acidic, returns to the nickel stripping circuit. The nickel is in cathode form for sale. (Dwg. NC-0011-SU, Figure 7.12) Cobalt, along with unextracted copper, nickel and zinc is recovered from the nickel liquid ion exchange raffinate by precipitation with ammonium hydro- sulfide, produced by sparging hydrogen sulfide into an excess of ammonia solu- tion. The sulfide precipitate is separated from the raffinate in a clarifier. The clarifier overflow is filtered of entrained precipitate and sent to ammonia recovery. The underflow is mixed with purges from copper-cobalt stripping of the liquid ion exchange reagent. The mixture is pressure leached with air to preferentially dissolve the nickel and cobalt sulfides, leaving copper and zinc sulfides undissolved in the residues. The latter are removed by filtra- tion and sold, as minor products, to smelters for recovery of metal value. Following pH adjustment and reprecipitation with hydrogen sulfide for final removal of any zinc and copper solubilized in the first leach, the nickel/cobalt sulfate solution is heated and autoclaved. Nickel is reduced with gaseous hydrogen. Sufficient ammonia solution is added during reduction to neutralize the acid formed. Only a portion of the nickel is removed per pass to prevent over-reduction and subsequent contamination of the nickel powder with cobalt. After densif ication through repeated recycle, the nickel powder is removed, washed, and passed to drying and briquetting for sale. The largely nickel-free cobalt sulfate solution passes to an evaporator /crystallizer , where the remaining nickel and cobalt are precipitated as the double salts with ammonium sulfate. Excess ammonium sulfate is purged, and the salts are redissolved in 7-14 ^ -^ 2 U ■i *• (kill 2 - u i" ■r iI k. n ^ <* > ^ V) h ^ K "T IV Hj ■i: ' ~ VI V- ;< ^ ^ .^ *0 .S 1 i 1 1 1 I 1 1 : 1 i MIM <}• ■ i i 1 1 r-H ' 1 ! ! 1 1 i 1 1 r^ i 1 i 1 ; 1 ' i ' 1 ■ ; ! 1 1 g ; ! 1 i S : ': ! 1 ; : 1 : ; ! 1 ^ :(0 Vi 1 i * ■ (^ • ' r^i 1 ;S!|&s8: ■,? (\j| >:, v^i . 1 fM 1 ^: 1 s 1^ ^ ^hi j^ !/^i.kJ , CM ■ 1 • CM J^ Ql i § 1 i CN, OM •^, [^ 1 1 ^ 5 1 y J- ^? ? h^ 5: si 5; ^ si !2 7-20 7-21 It has been assumed and documented that a slurry, whether produced from nodules or later ite ore, when leached at high temperature will substantially extract the metals of interest as well as undesirable trace metals in amounts depending on the source. The leach reactors are very fundamental in design and of proven operational reliability. A design change or equipment sub- stitution would not appreciably alter material and energy conservation, nor show significant operational savings for the capital invested. A possible alternative here is energy recovery from the leachate by indirect heat exchange. The consequences of this would be a need for more steam production and water utilization due to finite AT's, as well as a need for money and plant space investment in pumps and heat exchange equipment. The special alloy flash-down values used in this process are economically and mechanically feasible and present the most convenient arrangement. The copper raffinate neutralization step, which assumes the use of enough ammonia to generate a basic solution and oxidation of cobalt to prevent co-extraction with nickel, could be eliminated if a reagent could be found to selectively remove nickel in the presence of cobalt. The precipitation and separation of basic salts of iron, manganese, magnesium, and aluminum would be eliminated, with a savings of necessary equipment. The ammonium sulfate solutions purged from the liquid ion exchange/ex- traction and the copper raffinate neutralization have been treated with lime for the recovery of ammonia for recycle. An alternate approach would involve direct recovery of ammonium sulfate by evaporation/crystallization for sale as a byproduct. The alternative is highly energy intensive because of the number of crystallizing/evaporating steps necessary to produce a relatively pure byproduct from a stream containing many entrained process impurities. If this were done, 2100 tons of ammonium sulfate per day would be produced with a corresponding need for 420 tons of ammonia per day. There would be an 880 ton per day savings in lime utilization with a corresponding loss of 2140 tons per day of gypsum byproduct. The energy savings in this alternative are not significant. The production of ammonium sulfate is outside the interest of the process and could be severe on the local market, because the production is approximately 15% of present U.S. production. The proposed alternative of selectively extracting the nickel without cobalt oxidation would present a problem at the end of the process in the disposal of 1320 tons per day of soluble sulfates that would have been preci- pitated out. If direct disposal were not satisfactory, 1200 tons per day of additional limestone would be needed to form the gypsum-metal hydroxide precipitates. Alternative configurations of the metals separation steps are possible, such as selective extraction/selective stripping, but their impact on overall plant material and energy balances would be minor. Many variations are possible in the details of the scheme used for recovery of cobalt from a mijced sulfide precipitate. The impact on plant requirements would not differ appreciably from the present approach, since the sulfides will still be oxidized to sulfates and purged and the metals hydrogen reduced. It does not appear likely that the solution could be purified easily enough to permit recovery of electrolytic cobalt. 7-22 If environmental constraints, availability, or costs dictate that an alternative to the use of coal be found, the most likely substitute would be a residual type heavy fuel oil. A consumption of approximately 4,550 barrels/day would be required in place of the 1,080 tons/day coal shown in the balance. If the additional amount of process electrical power could not be purchased, on-site generation is possible. An additional 270 tons/day of coal (1,180 barrels/day fuel oil) would need to be purchased. A corresponding increase would be required in the amount of flue gas to be scrubbed for sulfur/ particulate removal (approximately 11%) and in plant cooling requirements (approximately 51%). To provide for water lost in wet cooling, an increase in make-up water of approximately 13% would be required. The use of dry cooling instead of conventional wet cooling towers would reduce water consumption at the expense of some increases in both capital costs and power requirements. Overall, the impact on plant energy requirements would be minor, but make-up water requirements would be reduced. 7.3 PLANT OPERATING SUMMARY TABLES The complete material and energy balances presented in the sectionalized flow sheets has been abstracted , and those streams which constitute plant inputs and outputs are summarized in Table 7.1. Included are production rates for the products, consumption rates for materials and supplies, raw materials requirements and a breakdown of the amounts of solid, liquid and gaseous wastes, and the requirements for fuels and purchased power. The stream numbers are keyed to the sectionalized flov; sheets. The distribution of services required within the plant has also been abstracted and is summarized in Table 7.2. The composition of those output streams which constitute process wastes is summarized in Tables 7.3 to 7.5. These compositions were derived from the overall plant material balance and are presented for only the major materials in the process. l\rhere trace amounts of a material are indicated, it has been assumed that previous removal steps will have reduced it to a level which is either acceptable for discharge directly or is present in amounts too low to quantify within the accuracy of the methods used in this study. A summary of the output of toxic substances in these waste streams is presented in Section 7.4. 7.4 POTENTIALLY TOXIC ELEMENT FLOWS AND DISTRIBUTION In the high temperature sulfuric acid leach process, the total input rate of inherently toxic elements is about 65 tons/day. Estimates of the distributions of the ten selected constituents were made for each process juncture by the procedure outlined in Section 4. These estimates or assumptions were then combined to produce a collection of estimated attenuation factors as displayed in Figure 7.16. A summary listing of estimated toxic element dis- tributions and flows to product and waste streams is presented in Table 7.6. 7-23 Table 7.1 Plant Inputs and Outputs High Temperature Sulfuric Acid Leach Process Stream No. Composition Rate Products 90 Cathode Copper 10^ tons/yr 32 150 Cathode Nickel 10^ tons/yr 38 181 Nickel Powder 10^ tons/yr 0.2 182 Cobalt Powder 103 tons/yr 4.1 178 Zinc/Copper Sulfide 10^ tons/yr 5.2 Ma.jor Inputs 220 Raw Nodules: Solids 10^ tons/yr 3,330 220 Transport Water 106 gal/yr 1,920 222 Coal 103 tons/yr 363 225 Limestone 10^ tons/yr 163 226 Lime 10^ tons/yr 380 14, 92, 126, 144, Sulfuric Acid 103 tons/yr 940 155, 177, 191 240 Water 10^ gal/yr 1,540 Purchased Power 10^ Kwh 148.9 Major Outputs 207 Tailings/Wastes: Solids 10^ tons/yr 4,220 207 Liquid 10^ gal/yr 1,120 221 Nodules Transport Water 1,920 248 Stacked Off-gases 10^ std cu ft/min 245 Low Level Emissions: 25, 193, 210, 224 Vents 10^ std cu ft/min 55.8 96, 160 Electrowinning Vents 10^ std cu ft/min 11.63 243 Cooling Tower Evaporation 213 Sanitary Wastes Production Materials and Supplies Gases NH3 103 tons/yr 2.4 H2S 10^ tons/yr 4.8 Ho 106 std cu ft/yr ^^-^co CI2 103 tons/yr ^-^^^ 144, 155, H2SO4 106 gal/yr 134.7 177 HNO3 103 gal/yr 2 Fuels (POL) 103 gal/yr 552 LIX Reagents 103 gal/yr 247 38 Flocculants 103 tons/yr 1 97 Electrowinning Additives tons/yr 16 153 Na2S04 tons/yr 610 154 H3BO3 tons/yr 120 152 Carbon tons/yr 40 252 Tower Water Treatment tons/yr 358 256 NaCl tons/yr 230 7-24 202 80, 174 180 252 Liquids 92, 126 191, 14 156 183, 228 Solid! 3 Table 7.2 High Temperature Sulfuric Acid Leach Process Services Profile Process Power Kw Ore Preparation 2,900 Leaching 1,350 Tailings Washing 600 pH Adjustment 450 Cu Extraction/Stripping 300 Cu Electrowinning 10,000 Neutralization 700 Ni Extraction/Stripp ing 600 Ni Electrowinning 16,650 Co Recovery 900 NH~ Recovery 750 Materials Handling 1,050 Services and General Plant Use 6,850 (24,300)*^^ Process Cooling Steam Water 10^ Ibs/hr Gal/min 383 (1) TOTAL 18,800 (130 ) 700 (3) 29,600 3,400 4.5 500 8.5 13 500 10.1 - 51.4 2,100 306 33,600 1,200 Process Water Gal/min 25 20 20 25 11 3,070 70,900(4) 3^200 (1) 3 81 X 10 Ibs/hr flashed steam exported- for process use, (2) (3) Generated internally. Used internally in power generation. ^^'*640 X 10^ BTU/hr 7-25 iH XI CO H tn w e en CC OJ cu o u o 4J u CD PM J-l T3 c •H QJ O 3 < iH m O M-l •H W V^ 3 T3 U-l •H ^ .H 3 O w cn 0) M-l ^ o 3 4J en nJ 4-1 Vj c q; 0) a 3 e 4-1 Q) •H H •W W X c 60 o •H u W H CO cn ■X3 T3 « X •H r-\ 00 3 iH ITI M-i )-i O CN U-i a CO O w c iH o •H X 4J CO ^ r^ -H o iH 3 W -a- • cri ro -^ < CN 00 «NO • s O c^ u o •H ,H 4J u CA) < W -N (n -I CO CO -- QJ T3 >-i C -H 0) -O "H 4J Vj O Cfl CO CAl IS X H CO ^ th -a 6 O i-H •H pa o tJ Crt c o X -H M W (1) cn 0) (U Ai CO OG C 3 V4 CO O 3 H 33 P-i o X O in CM CVI o O 00 cn o X 00 in o m CM TJ X 0) CO ^ .H vD O -H o 00 CO CO m 00 cu H J 4J c c o (U ^ CM & V4 in C/2 CO rH in CM 1^ m o CM o «3- o z s CO u C/3 O CM O r' ^ rH CO CO o ^^ H m CM u a. S H e 4-1 < u 3 CO CO (U i-l O O CMO fM • to CO O CM o o CO CO "H rH C 4J o o M <: s w CO m O CMO . U O ~ PQ V-i )-i 0) 3 1 tH CO hJ to iH •H to H >i M T3 M 0) 3 ,n rH CO WD to S s to Q) U u W) CX) e •H 0) Ul H O a j= e M -H CJ X o c (0 0) o 3 -o a) > pi O J-l « c ►J > CO CO to (U 4-1 o c o 0^ (U CO M CO % cfl H ,4= O X w >-l Xi o U 4-1 C (0 o a; 0) (0 > (0 w ^1 3 4-1 01 m Cu3= 3 vO a. n) as o ^ £ U P X rt w H tH O U U CO C 4-1 o CO u 3 -* CO CNI 4-1 .c CNJ CO >< 3 w p 00 in o o -a- o o as 1^ o o -3- 00 o ■* CO g g Pi H u-i i-( O r-~ CN r-H CS r^ Pi H o O O u-i O 00 00 as „ s ^ S 4-) (M d o o <; o z .. en e s •t • * a rt rH , CO ro C 3 c •H T3 3 t-l O X a a X a; (U 7-2J A cn Eh S X -l •H 4-1 14-1 QJ ^ M-i M-l 4-1 X QJ 0) O P 4-1 U rC o. •H 4-1 M c S cn iH •H ^ C O T3 o •<-* rQ 0) 0) o e ^ iH iH w >, 4-1 01 X> OJ m ^ iH O CO ^ P 0) 4-1 1—1 TD T3 ^ C O U !-i &0 •H CO ,-— N QJ C 4-1 X (n -H CO Cfl •H 0) X) -H s Xl C OJ 0) o C oj O 1-1 Vj QJ 60 QJ Vj O- C ^1 r^ c 4-1 .g #> CO QJ w e m iH QJ 4-1 P V- a X C 2 o cn OJ 4-1 cu p U x: S 4-1 CO 4-1 o •H • 14-1 iH 4-1 U) o MH cn U3 c 4-1 C 0) o CO c o •H U tn o o 4J c GJ t-i CO (U u a a P e o •H c OJ u X 4= 0) rH Cu o o 4-1 0) 4-1 CO 4J TJ 0) CO c 0) >. t-H QJ rH 1—1 -a > •H • iH TJ 0) •H CO cn CO -H 4-1 (3C 4-1 QJ •H O CO QJ U 4-1 CO e 0) TJ P c •H J= T3 (U O 4J ■u QJ QJ 4-1 -H cn ^ O O V-i 0) MH 4J O O- 3 O V4 y-i 0) O a U-t rH J= ^— s 4-1 o P 4-1 o cn Cfl o TD cn C OJ • 0) QJ O OJ u 1— 1 >. a •H >^ c GJ o 4-1 P II ^ S-i P 4J cn Ci ^ CO 4-1 u Cfl •H U c o M XI U QJ U o LD O O C c o ^ rH Xi C/5 H p^ o o o rH LD (Ti CO o (0 CO iH &1 cn fl. ^ CD h-1 > u < < CO M O O O o o o n \D c LO rH c r-\ rH o ^ iH Eh W Pi CD Pl4 O c/i H tJ 1=1 H I— I H CO Z o o rJ 7-29 > U <: U W Eh CD O in r-H rH O O -^ (T< -^ r-i U D^ Ti en XI nH > u < u < CO H in (Ti iH in ry\ 0^ rM O O ■^ 1-1 o o O I TI tn < r-i EH O o o o (Ti in (N H > U u en < in o 01 o o Ol (0 S-i Cn T) W, J3 iH iJ>0 u fit Ui ^r^ iH in m on rH o CN r- o 00 tH • • • • • H Z O u o in Eh W K In w o '3' ^ Ol fd rH X > TO DIS .25 L .25 T .10 P 7-30 U Cn Tl XI t-H u -a 43 rH > U < o w EH > u u W Eh o O o CN O 00 u < U > U u o in in o W iH ^ o in O O H t-H o o o O O ^ O o o o t t rn • o J EH rt C/1 :n o a p< H Ul M H () U w Tl X} u w s in ■* o o o M •d- o O EH ■ H U a 2 u r-t h^l Eh > C) H O X >H W n r- U ^.0 (N in §s n o C5 o o o o H c n a n 3 O H EH I ) H) U) XI s u pi; U) H H f 3 X U) iH t^J 3 W 00 00 (^ H o o o -X ■ • 1 rt s l-l H Eh Ci< U C) d- W H O hJ u CP U H a " u < U k X r^ no Lf> ^ C) 1^ o '^ tl O o o u en Tl XI iH > V in <; o w H -* iH O) ^ o in M ^r in o o -* oo o o M Tl rH X! r-H > in W w H > w 1^ 1.0 ■* va iH ■=T in G o m m in o n o o o o o o ro o M Tl X) -H CJ U W Eh ^ o o^ vo in o o o m o Q P M o CJ w PS § 7-31 I »J W S ja r-\ o w ErI (ii n w m n c o CM N O o Tl C (d > -a o w 3 in O U <* o 0^ O ■a x> H K o o o u r- Eh H 00 o (N r CO I S X O H 1-1 H ^ s m u P C5 H Q u 2 a 1 XI rH > w H VD VD ro fNJ O o C o o M T) ,Q rH > u in U o CO H kD ^ a\ n ■ii 00 c^ o n W H o r^ ^ LD C5 ro C3 O m O w o w Q t Table 7.6 Summary Flows and Distribution of Potentially Toxic Elements High Temperature Sulfuric Acid Leach Process Input Rate Distri to Plant Report Element 10^ lbs/day Waste Barium 76.2 100 Lanthanum 31.4 50 Vanad ium 7.6 11 Chromium 0.19 90 Silver 0.060 10 Cadmium 0.38 5 Arsenic 0.96 11 Antimony 0.58 11 Thallium 3.8 51 Lead 9.6 90 rcent Discharge Rate to Waste oduct 103 lbs/day 76.2 50 15.7 89 0.84 10 0.17 90 0.006 95 0.019 89 0.106 89 0.064 A9 1.94 10 8.64 The outputs from the high temperature sulfuric acid leach process include gaseous emissions, liquid waste streams and solid waste streams. Liquid and solid wastes are, in some cases, combined in a single waste stream as a slurry or slime. The potentially toxic elements are distributed in varying amounts in certain of these streams. The waste streams are listed in Table 7.7 with the total flow, in tons/ day, representing the sum of the liquid, solid and toxic element components of that stream. Streams 77 and 152 will be incinerated with other combustibles, with any residue going to the tailings pond. Stream 251 will contain the material which must be removed from the off-gases to permit discharge to the atmosphere, and may contain significant amounts of hazardous substances. Streams 91, 151 and 213 will contain one or more of the potentially toxic elements. While it is possible that other process waste streams will also contain hazardous substances, such as water treatment chemicals (streams 244, 250) or combusion ash constituents (streams 247, 251), these have not been considered here for reasons explained in Section 4. In addition to the total flow rate and distributions of potentially toxic elements in the process streams, it is important to define their concentrations as well. This can be done by using the flow rates calculated from the atten- uations factors summarized in Figure 7.16 and the total flow rate of the streams calculated from the process material balance. The concentration of hazardous substances in the waste streams from the high temperature sulfuric acid leach process is summarized in Table 7.8. 7-33 H •V o 3 !«! y o c c •H (X M O 8 (U M • 4-» O •a •o T3 ■T3 x) T) •T3 T) xt T3 c C C a C d c (3 C C C C o O o o o o O O O o O PL. Pm •H U PUl CM •H 4-1 p-1 CM CM CM CM Oj to M (0 to M n) m in CO CO CO CO &0 60 M 00 60 M 60 60 60 60 60 60 c c (U c C 0) c C C c C C 1-1 •H c •H •H c •H •H •H •H •H •H I-l r-( •H iH rH •H f-l r-* iH tH iH ■H •H •H o •H •H o •H •H •H •H •H •H cd n) Ci (d CO c cfl rt to tfl Cfl CO H H M H H I-l H H H H H H u 60 U CM o •> > H CO ^ hJ CO CO CO CQ < T3 U :» /-N o >1 CO b -o CX3 rH CO CI (0 C 0^ u o •« O 4J O H ^— ' CN o CM O CM d o o o\ iri m rH r~- O -* >* m CM CSJ CM rH c • . 4= CO •a 60 60 60 4-1 m 60 60 •H U O to •rt 3 V-l to MH CO CO CO o rH C to c: u CO •H U 3 3 Q CO tb u T) o CO :3 to u O a) U V y Corporation 1 1 >- -' ■4 ^'^^ z>ocU r»- o "^ to iSO in 6 o u Z ^ -^' s^ ,tS,l !/ u/? M ^ll IJ ^1 c-1 U II 1 1 5-2 much lower acid concentrations are appropriate In the case of nickel, which also is produced as cathodes for sale. Of the several options for manganese recovery from nodules, that selected for the base-case process involves drying of the final aqueous chloride solution to produce a dry, impure manganese chloride. The manganese chloride is charged to a high temperature electrolysis furnace, where it dissolves in a molten alkali chloride bath. Electrolysis of the bath liberates molten manganese, which is tapped from the furnace, cast into molds and sold as manganese metal. Fused salt impurities are also skimmed off, solidified and sent to waste disposal. The second major product of the electrolysis is chlorine gas, which would be recovered along with the chlorine produced in the initial hydrochlor- ination step. Recovered chlorine would need to be reconverted to hydrogen chloride or otherwise utilized off -site in a large scale chemical process, such as the manufacture of polyvinyl chloride. Plant services include facilities for raising the necessary steam and part of the power required for process use, supplying the make-up water and cooling required, and providing for materials handling for process materials and supplies. While data on the reduction of nodules with hydrogen chloride and subse- quent recovery of metal values from acid chloride solution have been reported in the patent literature, this technology currently has no direct analog in commercial extractive metallurgy. Also, while the thermodynamics of the re- duction step are well known, and reports on laboratory studies of some separa- tions of metals fronj chloride solutions are available, no data exist to indicate the expected properties of the nodule residues (e.g., filtering rates) or the problems associated with recovering a salable manganese metal from impure chloride solutions. Thus, the proposed process configuration has been established based on literature data which contain little detailed engineering and design information and without the benefit of insights which might be gained from analogies to conventional technology in related operations. This means that alternative process configurations might well be technically and economically more attractive, but access to more data would be required to make such a judgment. One alternative configuration would involve the reduction of nodules by strong aqueous hydrochloric acid, thus eliminating the costs of drying nodules for high temperature roasting. However, iron would then be dissolved, and its removal would require additional and probably costly process steps. Alter- nating processes for the separation and purification of copper, nickel and cobalt are possible but would have little impact on overall process inputs and outputs. Alternative methods for manganese recovery would, however, have major impacts since it is the major metal produced. Direct electrodeposition from impure chloride solution, if possible, would simplify the process scheme. Precipitation on aluminum, i.e., cementation, has also been suggested and, while probably reducing energy requirements, it would require additional process steps and involvement with the scrap aluminum/alumina markets. 8-3 A key process alternative, and one which will impact siting requirements considerably, pertains to the arrangement for returning hydrogen chloride to the process by exchange for or regeneration from the chlorine evolved. While a buy-back arrangement with an adjacent hydrochlorination facility has been assumed, hydrogen chloride could be regenerated on-site by burning evolved and purchased make-up chlorine with purchased hydrogen, or hydrogen formed from purchased hydrocarbons or coal. Obviously, each alternative would have a distinctly different set of inputs and outputs, but the operation would be on a large scale for any of the options mentioned. 8.1.2 Detailed Process Description The hydrochloric acid process is a four-metal process in which manganese, copper, nickel, cobalt are liberated from dried nodules by a high temperature (500°C) gaseous hydrogen chloride treatment of nodules. Hydrogen chloride reduces manganese dioxide to manganous chloride (liberating chlorine gas) and also reacts with other metal oxides to form soluble chloride salts. A hydro- lysis reaction and quench follow, where water is sprayed on the nodules and the iron is precipitated as ferric hydroxide. The nodules are leached v/ith aqueous hydrochloric acid, forming a concentrated pregnant liquor of chloride salts. Copper is extracted by liquid ion exchange reagents from the pregnant liquor, stripped and recovered as electrowon cathodes. Cobalt is solvent- extracted from the copper raffinate, stripped, and separated by precipitation with hydrogen sulfide and is recovered from the sulfide precipitate by select- ive leaching and hydrogen reduction, along with some nickel, zinc and copper. Nickel is extracted by liquid ion exchange reagents from the cobalt raffinate, stripped and recovered as electrolytic cathodes. The nickel raffinate is evaporated, crystallizing manganese chloride as well as the other remaining chloride salts. The salts are dried using combustion gases in a counter current dryer. The dried salts are charged to a high temperature fused salts electro- lysis furnace, where molten manganese metal is tapped and cast as product and chloride gas is liberated. Excess hydrogen chloride gas in the process is recovered and recycled. Generated chlorine gas is recovered, dried and deliver- ed to a local chemical complex which, in exchange, returns make-up hydrogen chloride to the process. Detailed descriptions of each segment of the process are given in the following flow sheets. (Dwg.*NC-0001-CL, Figure 8.2) Wet nodules are reclaimed from storage and fed through a primary cage mill where they are reduced to -7/8". They then pass to a fluid-bed dryer where surface and pore water is removed at 175°C by direct contact drying with com- bustion gases. Bed overflow is reduced to -200 mesh in a secondary cage mill, and dryer and mill fines are removed from off-gases with cyclones and an electro- static precipitator. Off-gases pass to gas treatment for scrubbing, while dried nodules are transferred hot in an enclosed conveyor to reduction. (Dwg. NC-0002-CL. Figure 8.3) Dried nodules are reacted in a fluidized bed hydrochlorination reactor by contact with hydrogen chloride gas from the HCl surge, which is preheated to 175°C. The exothermic reactions maintain the reactor at 500°C. Manganese dioxide is reduced to manganous chloride and chlorine. Essentially all the copper, nickel and cobalt, as well as 90% of the alkali and alkaline earths 8-4 <^ oil I B li . : C^ • 1 00 1 w 1 oi 1 :=) : 1 o M 1 [x< j«' -^ Jo > UJ iD U J o I o o I o J-5 8-6 present in the nodules, react to form chloride salts. Approximately 100% excess HCl gas is used in the reaction. This HCl gas, as well as the chlorine and water liberated in the reaction are returned to HCI/CI2 recovery. The reduced-chlorinated nodules are delivered to a second fluidized bed where any iron chloride is hydrolyzed by a water-spray quench, forming insoluble ferric hydroxide, and the nodules are cooled to 200°C. The off-gases from both fluidized beds are sent to a cyclone-electrostatic precipitators system for dust removal. The hydrolysis off-gas is also passed through a waste heat recovery system. (Dwg. NC-0003-CL, Figure 8.4) The chlorinated nodules are leached with water and HCl in a tank where the soluble chlorides are dissolved to form a liquor with a pH of 2. The heat of solution is removed and the solution is cooled to 40°C by circulating the liquor through an external heat exchanger. The slurried nodules are washed counter currently in a 6-stage thickener circuit to remove 98% of all soluble metal values, forming a pregnant liquor. The wash water is recycled water from process water surge. Wash tails are sent to the waste treatment area. Floccu- lant is added to the thickeners to improve the settling properties of the solids . (Dwg. NC-0004-CL, Figure 8.5) Pregnant liquor is filtered and passed to a three-stage counter current liquid ion exchange extraction circuit, where copper is removed from pregnant liquor. The preferred liquid ion exchange reagent is LIX-65N. Sodium hydroxide solution is added at interstages to keep the pH of the pregnant liquor constant at 2, to insure complete extraction of the copper. Some of the other chloride salts are physically entrained in the organic phase. These latter chlorides are removed by washing with a water solution in a two-stage wash circuit. A purge from this wash solution is returned to the pregnant liquor surge, and make-up water is added to the circuit equal to this purge. The copper is stripped from the organic by counter current contact in two stages at controlled pH with depleted (weak) electrolyte from copper electro- winning. Since electrowinning conventionally occurs at a higher temperature than the operation of the liquid ion exchange circuit, the strip solution is heated before passing finally to electrowinning. Provision is made for cleaning the mrxer-settler units used in extraction and stripping periodically and for recovering organic and aqueous phases for recycle. Degraded organic, dust, and other forms of "crud" are removed from the organic and incinerated. Since a small amount of cobalt is co-extracted and is not stripped with the copper, it must be removed from the liquid ion exchange reagent to prevent its build-up. This is accomplished by precipitating the cobalt from a purge stream of organic with hydrogen sulfide. The precipi- tated solids are washed from the organic and passed to cobalt recovery, while the purified organic is returned to the extraction loop. Make-up organic is added to the stripped reagent to offset degradation and soluble organic losses to pregnant liquor and water wash, and the metal-free organic is recycled to extraction. o- 7 k c v:r g ^ -J ^ 1 o 1' (PJ \ i 1 2^ ?^ 1 Vj 5 *^^ . , 1 — ^ J^ / ^ 3 M 00 ^" ^ ^ (A) — ~ -— — O M rr- r- r-"' P4 ( \) ^ ■^ $ 1 rs^ Vn >? ' ? 1 $ y v'' n - il o >> i^\-~ ^ o - <\ '^??^ r ?^ "^ ,^^!i ^.^ ^ ' V . 1* S^ - ^ — y - O k s, \ - 1^ <4 Vi A vi n - , 1 ^■1 8> ^ O ^ >}■ r in n S \ \ Is 'i ^ Z ft. 1 ?i i 1 J' \ W k < V, ^j vy 8-9 (Dwg. NC-0005-CL, Figure 8.6) Cathode copper is recovered from the strong electrolyte from liquid ion exchange using conventional technology. Starter sheets are deposited on titanium blanks from strong electrolyte in the stripper section, are removed, washed, looped, and returned to the commercial section as starters. Full term cathodes produced in the commercial section are washed, unloaded, and prepared for shipment as cathode to sale. The major part of the weak electro- lyte is recycled to the liquid ion exchange section for stripping, but since a small amount of nickel is co-stripped with the copper and not deposited in electrowinning, it must be purged from the system. The purged electrolyte passes through purification cells where copper is removed by electrowinning to extinction. The decopperized electrolyte is sent to cobalt recovery for further processing. Make-up acid is used to redissolve scrap copper for return to the commercial cells for deposition and sufficient steam, wash water, and make-up water are added to the circuit to offset water which vaporizes and is carried with evolved oxygen, from the electrowinning cells. (Dwg. NC-0006-CL, Figure 8.7) The raffinate from copper extraction is adjusted to a pH of 4 by the addition of NaOH solution in a surge tank. This stream then passes to a two-stage counter current solvent extraction circuit where cobalt is extracted as a tetrachloro-complex by the organic reagent. The preferred organic reagent is a tertiary amine such as tri-isooctylamine. The cobalt is stripped from the organic by counter current extraction contact in two stages with recycled process water. Because of the high manganese concentration in the raffinate, a portion of it is extracted with the cobalt. The cobalt is separated by addition of hydrogen sulfide, which reacts to form cobalt sulfide precipitate which is centrifuged from the liquor and sent to cobalt recovery. The liquor, which is basically manganese chloride solution, is sent to manganese recovery. (Dwg. NC-0007-CL, Figure 8.8) Nickel is recovered in a manner similar to that used in copper extraction. The cobalt-free raffinate passes to a three-stage counter current liquid ion exchange extraction circuit where nickel is removed from the raffinate. Sodium hydroxide solution is added at interstages to keep the pH at 4 to insure good nickel extraction. The preferred liquid ion exchange reagent is LIX 65N. The loaded organic is washed to remove chloride in a two-stage operation. The nickel is stripped from the organic by counter current contact in two stages with depleted (weak) electrolyte from nickel electrowinning. Similarly to the copper circuit, heat is exchanged between the incoming and outgoing streams. Provisions are also made for cleaning the mixer-settler units and sending the resultant "crud" to incineration. An organic purge is also taken to insure against build-up of unstrippable material on the organic. Make-up organic is added to the stripped reagent to offset degradation and soluble organic losses. 8-10 f^ - fc -^ "^ it *n 1^ In -k f5! 8 M CD :^ 8 ^ rvi r f ^ I" c «« c SS S « $1 r- I 2 111 3 T J\ ■ 6 oH 5-11 5> a a 1 i ^ ( 1 5 5 1 "1 O' -O] S § 8^: s? ol -.1 ^ 2 xS o CT ^ "Ml OJ _^ ' o ■ \- ' 3 ^ 7 . \0 :2 _1 u in I VQ > ^ o ^ R (\j VI ~^ B;'^ -- ::^s "o I ?0 Vn iTl iO "^ ^ N ^ « ^ N ^ $ fQ fO O \ • 2 5 Kl -^ 58 '■ ^jQ-!;-: . ■■ < 55 !c V. o ^ . «M ^ V> . 1 5 fccv- "^S.: -^ 1 O 1/1' j> i i?) -: 1 C!^ f^" ^ 1 1' mm 00 W Pi O 5-15 5 1 $ § < 51 ^ (XJ f ■^ 5 ^ ^1 ^ ^ - <0 ^-vi § ^ r\4 ^ i : ^ .; 5 XI 'Vi § ^ y CM r- 6 ^^^' fi 1^ 6 1 ^ 1 ti s^ 8 w Pi o 8-16 8-17 recovery of metal value. Following pH adjustment and reprecipitation with hydrogen sulfide for final removal of any zinc and copper solubilized in the first leach, the nickel/cobalt sulfate solution is heated and autoclaved, and nickel is reduced with gaseous hydrogen. Sufficient ammonia solutions are added during reduction to neutralize the acid formed. Only a portion of the nickel is removed per pass, .to prevent over reduction and subsequent contam- ination of the nickel powder with cobalt. After densif ication through repeated recycle, the nickel powder is removed, washed, and passes to drying and briquetting for sale. The largely nickel-free cobalt sulfate solution passes to an evaporator/crystallizer , where the remaining nickel and cobalt are precipitated as the double salts with ammonium sulfate. Excess ammonium sulfate is purged, and the salts are redissolved in strong ammonia solution. The cobalt in solution is oxidized to the cobaltic state with air. This permits the cobalt to remain in solution when the stream is subsequently acidified to remove nickel salts, which are then separated and recycled to the pH adjustment step. The nickel-free solution is then heated and autoclaved for removal of cobalt by hydrogen reduction. Sufficient ammonia is added to neutralize the acid generated. The cobalt powder is dried and briquetted for sale, with the ammonium sulfate purged to the lime boil. (Dwg. NC-0011-CL, Figure 8.12) The off-gas from the hydrochlorination reactor contains a mixture of unreacted HCl , chlorine and water. The HCl is absorbed by water, forming a highly concentrated HCl acid solution. The overhead gas from the tower is combined with the off-gas from electrolysis and is dried by passing through a sulfuric acid solution, which strongly absorbs the water from the gas, leaving dry chlorine for delivery as product. The HCl gas from the hydrolysis reactor is absorbed in the second tower, producing strong hydrochloric acid solution. The off -gas from this tower is mainly humid air and is sent to gas treatment. In both absorption towers, the dissolution of HCl is highly exothermic, and the heat is removed with cooling water. A series of lean HCl vents is scrubbed to remove the last remaining HCl in a third tower. The overhead gas from this tower is also sent to gas treat- ment. All the HCl acid solutions are combined with sulfuric acid, and the HCl is stripped and sent to a hydrogen chloride surge, where make-up hydrogen chloride is added, and sent to the hydrochlorination reactor. The bottoms from the HCl stripper are sent to a water stripper, where water is taken over- head and condensed and the bottoms are sulfuric acid. The sulfuric acid is filtered to remove any particulate matter which may have been entrained in the gas streams entering the HCl recovery section. The resulting strong H2SO4 is recycled. An aqueous HCl stream is split from the main aqueous HCl stream to provide an acid stream needed at various stages in the process. Excess water developed in HCl recovery is returned to the process water surge. (Dwg. NC-0012-CL, Figure 8.13) Ammonia is recovered from the ammonia sulfate purges from cobalt recovery by reaction with slaked lime. Steam is blown into the mixture to strip the evolved ammonia. The vapor is combined with other ammonia vents, and they are scrubbed with water to form aqueous ammonium solution which is returned to the aqueous ammonia storage for recycle within the process. The gypsum slurry from 8-18 8-19 I> 1 J Vi 5? $ ^> ■4 q 1'' # y^ ^ ^ ^ < ^5 J| U I 1 m fNj 0) I o w oi o M M ^ n 8-20 the lime boil is cooled and combined with the slurry from stack gas treatment. After liquid/solid separation in a thickener, the solids are combined with other process solid and liquid wastes and plant run-off, heated for pH control, as required, and pumped to the tailings impoundment area for disposal. The overflow from the thickener is returned to the process water surge, where it is combined with other water returns and make-up water to form the process water needs, as required. (Dwg. NC-0013-CL, Figure 8.14) Facilities are provided within the plant for receiving and reclaiming raw nodules, coal, lime and limestone, hydrogen chloride and chlorine, ammonia, and other process materials and fuel. (Dwg. NC-0014-CL, Figure 8.15) Plant services include process and cooling water supply and treatment, steam raising and power generation, gas treatment and combustion gas preparation. Make-up water is clarified and softened for distribution to the process, as required. Additional treatment is required for cooling tower water make-up, and for supplying plant potable water. Process combustible wastes are burned, along with additional coal, in the main boilers to raise the required process steam and generate a portion of the power required in the process. Following particulate removal, the flue gases are combined with other process off-gases and pass to gas treatment, where sulfur oxides and other acidic constituents are removed by scrubbing with limestone. The scrubbed off-gases are combined with scrubbed vents from ammonia recovery and stacked. 8.2 PROCESS ALTERNATIVES A major processing alternative is to use a low temperature aqueous reduction/hydrochloric acid leach. The stoichiometry of the reactions would consume the same amount of HCl as the high temperature gaseous reduction. However, the iron would be dissolved as ferric chloride and would necessitate more expensive solvent extraction steps to remove the iron before recovering the valuable metal products and a spray roast of the resulting iron chloride solution to recover the hydrogen chloride. On the positive side, an aqueous process would not require the drying step. From a practical point of view, the best pilot plant data reported in the literature were on the high tempera- ture gaseous process. A number of possible alternatives are available for recovering manganese. Direct electrowinning of manganese has been proposed, but it is questionable that the final pregnant liquor would be pure enough to develop a good electro- won product. A cementation on aluminum to obtain manganese has also been proposed. The resulting aluminum chloride would be spray roasted to recover hydrogen chloride and form aluminum oxide as a product, which must be sold or disposed of. In addition, the manganese product may be contaminated with aluminum. Other schemes propose producing manganese hydroxide as a product which, in effect, sidesteps the problem of obtaining a pure metal product. A number of metal separation schemes are possible. These include using different organic reagents as solvent extractants or liquid ion exchange reagents. Some of these reagents, such as the Kelex series, may be just as 8-21 1 1 1 1 1 _i. ! , ji : ' MM'! ■ i : i i ill 1 ;5 CM - ■ 1 i I S fNJ 8 ^ W1 -^ 06 fv "1 -5 ' ■ :^^l 5' ;§} : I .^ I !« ' Iv ■ Q v » W Pi :=> o M 8-22 5-23 appropriate as our choices as suggested in the patent literature, whereas others are even further from commercialization than the proposed reagents. One potential problem of this process is the amount of chlorine gas produced. The amount of this gas (555 X 10^ tons/yr) is larger than the production of any chlorine plant in the world, so that the assumed neighboring chemical complex must be massive. The alternative might be to buy hydrogen and burn it with chlorine to produce hydrogen chloride directly. However, the amount of hydrogen needed for this process represents 8.5% of the total U.S. production of hydrogen for 1975, so that a "sell" problem would only be converted into a "buy" problem. The final alternative, of course, is to construct a plant with a smaller nodules processing capacity to match it more closely to existing caustic chlorine/hydrochlorination complexes. If environmental constraints, availability, or costs dictated that an alternative to the use of coal be found, the most likely substitute would be a residual type heavy fuel oil. A consumption of approximately 5,525 barrels/day would be required in place of the 433 X 10 tons/yr shown in the balance. The use of dry cooling in place of conventional wet cooling towers would reduce water consumption at the expense of some increases in both capital costs and power requirements. Overall, the impact on plant energy requirements would be minor, but make-up water requirements would be reduced by more than 73% to 37,000 gallons/hr. If it were not possible to purchase the required amount of additional process power, it could be generated on-site. This would require the purchase of an additional 1290 tons/day coal (5,650 barrels/day fuel oil). A corres- ponding increase would be required in the amount of flue gas to be scrubbed for sulfur/particulate removal, approximately 60%, and in-plant cooling requirements, approximately 120% in this case. This, in turn, would increase make-up water requirements by approximately 100% to provide for the water lost in wet cooling. 8.3 PLANT OPERATING SUMMARY TABLES The complete material and energy balances presented in the sectionalized flow sheets have been abstracted and those streams which constitute plant inputs and outputs are summarized in Table 8.1. Included are production rates for the products, consumption rates for materials and supplies, raw materials requirements and a breakdown of the amounts of solid, liquid and gaseous wastes, and the requirements for fuels and purchased power. The stream numbers are keyed to the sectionalized flow sheets. The distribution of services required within the plant has also been abstracted and is summarized in Table 8.2. 8-2A Table 8.1 Plant Inputs and Outputs Reduction/Hydrochloric Acid Leach Process Stream No . Composition Rate Products 73 Cathode Copper 103 tons/yr 9.6 142 Cathode Nickel 10^ tons/yr 12 195 Nickel Powder 103 tons/yr 0.2 196 Cobalt Powder 10^ tons/yr 1.9 190 Zinc/Copper Sulfide 10 tons/yr 1.6 163 I'langanese Metal 10-^ tons/yr 206.8 Major Inputs 260 Raw Nodules: Solids 10^ tons/yr 1238 260 Transport Water 10^ gal/yr 713 224 Hydrogen Chloride 10-^ tons/yr 724 262 Coal 103 tons/yr 459 266 Limestone 10-^ tons/yr 15.5 267 Lime 10^ tons/yr 4.2 52, 91, 112 50% Caustic Solution (as NaOH) 10^ tons/yr 29.7 280 Water 10 gal/yr 1100 Purchased Power 106 Kwh 756 Major Outputs 243 Tailings/Wastes: Solids 103 tons/yr 334 243 Liquids 10^ gal/yr 220 173 Fused Salts 103 tons/yr 233 222 Chlorine 10^ tons/yr 555 261 Nodules Transport Water 106 gal/yr 570 300 Stacked Off-gases 10^ std cu ft/min 565 Low Level Emissions : 205, 265 Vents 10^ std cu ft/min 52.3 77, 150 Electrowinning Vents 103 std cu ft/min 3.46 284 Cooling Tower Evaporation 10^ std cu ft/min 266 Sanitary Wastes Production Materials and Supplies Gases 64, 99, 124, H3S 103 tons/yr 1.8 143, 174, 206 245 NH3 tons/yr 50 175 N2 tons/yr 187 192 H2 106 std cu ft/yr 41 287 CI2 10^ tons/yr 0.1 (CONTINUED) 3-25 Table 8.1 (Continued) Stream No. Composition R.ate Production Materials and Supplies (cont'd) Liquids 76, 147, 194, 232 H2SO4 10^ gal/yr 430 149 HNO3 103 gal/yr 1 197, 270 Fuels (POL) 10^ gal/yr 500 53, 92, 113 LIX Reagents 103 gal/yr 57 Solids 37 Flocculants tons/yr 80 75 Electrowinning Additives tons/yr 5 145 Na2S04 tons/yr 360 146 H3BO3 tons/yr 52 148 Carbon tons/yr 20 176 Borax tons/jrr 67 182 Electrodes tons/yr 400 282 Tower Water Treatment tons/yr 440 289 Boiler Water Treatment, NaCl tons/yr 60 The composition of those output streams which constitute process wastes is summarized in Tables 8.3 to 8.5. These compositions were derived from the overall plant material balance and are presented for only the major materials in the process. Where trace amounts of a material are indicated, it has been assumed that previous removal steps will have reduced it to a level which is either acceptable for discharge directly, or it is present in amounts too low to quantify within the accuracy of the methods used in this study. A summary of the output of toxic substances in these waste streams in presented in Section 8.4. 8.4 POTENTIALLY TOXIC ELEMENT FLOWS AND DISTRIBUTIONS In the Reduction/hydrochloric acid leach process, the total input rate of inherently toxic elements is about 19 tons/day. Estimates of the distri- butions of the ten selected constitiuents were made for each process juncture by the procedure outlined in Section 4. These estimates or assumptions were then combined to produce a collection of estimated attenuation factors as displayed in Figure 8.16. A summary listing of estimated toxic element dis- tributions and flows to product and waste streams is presented in Table 8.6. 8-26 Table 8.2 Reduction/Hydrochloric Acid Leach Process Services Profile Ore Preparation and Drying Hydrochlorination Leaching /Washing Liquid Ion Exchanges Copper Electrowinning Nickel Electrowinning Manganese Recovery Cobalt Recovery Hydrogen Chloride Recovery Waste Recovery Materials Handling Services a^id General Plant Use TOTAL Process Power Kw Process Steam 10^ Ibs/hr Cooling Water gal/min Process Water gal/min 1,100 - - - 1,700 16.9 6,800 11.8 300 - 6,900 2.1 400 3.7 360 29 3,300 2.6 - 7.6 5,600 3.3 - 7.5 100,750 107.6 5,200 3.6 350 8.6 1,000 - 1,800 660 72,200 - 100 5.5 210 120 650 - - - 7,950 (29,50o{^^ (190)^^-* 810 1,500 2120 94,500 94,200 2300 (1) (2) Generated internally. Used internally in power generation, ^"^^850 X 10^ BTU/hr, 3-27 Cfi CO CO (1) e a nj o 01 u S-i Ph 4J w ji; o 4-J cd C 0) 0) i-J 3 iH ^3 M-l •H M-l O W < -d O •H ■H rH U O O CO tH 4= M-l O o O u en -o 4-1 > C 3: cu --^ 3 c 4-1 •H •H 4-J 4-1 U5 CJ C 3 o -o U (U oi -a CO (U 4-1 (0 tH 3 CO fe m CO TD -i s »J U !-i QJ CX X) a X 3 1— t M u ^ CJ rH 0) (U QJ ^ j^ CO bD (J c 3 U •H nJ 3 Z H K (i4 S-l CU cu QJ ex ^ CO M cu C 3 S-i ct3 3 H ffi Cu tH *^^ -3 CU J^ (0 iH CO •H CU CO ^ e CO QJ >-< 4-1 CO in u CJi u en t-H CO CO 3 f>n tN ■z u s .-=-. 1— 1 r-) CN OJ 3 hJ r-i U-t XI •H ffl •u -^ •H C w o O -H &. -P O O 3 O T3 dJ Pi M 0) N . dJ .H J-i e iH U •H O 3 hJ PQ .-( cn 4-1 c <-3 Q) >- cn e J nj ■U ao M C rt ^3 3 0) 3 ^ u r-l C/: H c/3 (U o. cn >. 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H A 13 CO 43 rH 43 (_3 tj) <: O O rH ty. rH (T. CJN 00 o o CO H Q CO £5 o CO H w !=> H M H CO c u x o H < M H H o (Vl o o M H P3 P3 M es H ca c« w u o Pi PM sc <: w < PS o o M H i Pi 8-31 \ > O en H u M -O ,Q rH J3 )-l T3 tn ^ .-H ^ > o u < CO H PL. 0^ a^ a^ c^ o O ON LTl U~l o a^ in o CTs 00 O O CT^ o> m o ^ o LO 00 -H ro 00 Pi r- O o o > 2 • • O o O M W H H cs hj <; ^ ^ < iri m 60 13 H-<-ON- PQ M CNJ <: U ^ O PL, K <: U M ^ O PQ O S iH o w O M o pi O 00 o PL, H >^ M > d o ro 0^ r-- O rH ^ O O H ft rH .H 00 -* O w ,a ■H ^ < CO H pL, o CNl O r- CO (^ -* O W o o O O ►J >-i • • • • CJ H Pi >< p hJ w U g <1 > u bO T) w <; M o > u <; o Pi O CJ c; w o- ON O c^ VD 00 00 ~d- H r^ o O 00 o o o o 8-32 03 m PQ kJ > U M) T3 03 ^ rH ^ U < U < c/) H Ph ro 00 .— I rH 00 OOv£)r-OCOvO^ ro 00 O rn ro O O r^ r-- O W & Z H H Z O O > PQ (J > O o o m 00 CTi CTi r~- o O c ON ?-33 O o ^ -H m H iH 00 o w VO O H vD O ^ W Q W CD N] > <: M Pi .H tN O CM <• Ph on ^ < O O Crt tj cd cS H <3 CO hJ > t-J w < o 0-) CO P-, o o ro C/3 CTs a^ O o M • • • W O V3 B o H CO n) tn ^ rH pq hJ > < c/: H vX3 ^D 00 o o vD ^ '^O o a> cT\ o o 'X) o ^ iH ^ w >^ W H pL, <; PJ w r-\ > ro ro 00 CN o r-^ r-~ O O o O O O vD s * * ■ ' H ^ u bOT > u < CJ PQ O .H U ro 00 iH o r~ O 00 O VO O O O H • p w M H S O a Pi o ON 8-34 I ij w w z CJ o M h-l Z H •~-^ u H & i-l Q <1 W M OS O U P^ M g o d- t/1 c M M-a cn X> ■H rCl ISJ > O <: o < to H PL. P ON 00 J--' in iH 00 vi- ON <; w H rH 00 M Op'O M ^ iH JO vO vO 00 ~3- ON 00 -* rOfno-*i-lcsroo OOOOvONDOO OOOOOOOO S M M 60 XI W > U H a o CJ Pi ^s^ ;::j cyj ac ■--. o en U H o W -< Pi en ^ rH M <; CO H § m (J CM ro «* s r-~ r>. o O O o o o o H < PP o )-l T3 CJ > u U o in H in -a- -a- NO o o o o o m rH + NO ON + CM 1^ CJ) z M hJ z o 5 M O Z Pi H CJ w ij u i k en 43 rH <: en H ^ NO VO CM CM en ro CO O o O o O ^ (Xl O O o p< en M ON Q o o •a- o )-i -0 o ~a- rH -►H > (J CJ o en o in Q vO ~3- -3- M n O O hJ O O o o O O o en 3-35 Table 8.6 Summary Flows and Distribution of Potentially Toxic Elements Reduction/Hydrochloric Acid Leach Process Element Input Rate Distr To Plant Repor 103 lbs/day Waste 22.0 100 8.8 100 2.2 14 0.066 90 0.018 90 0.11 1 0.29 11 0.17 11 1.15 51 2.9 90 Distribution as Percent g to: Product Discharge Rate to Waste 1Q3 lbs/day Barium Lanthanum Vanadium Chromium Silver Cadmium Arsenic Antimony Thallium Lead 86 10 10 99 89 89 49 10 22.0 8.8 0.31 0.059 0.016 0.001 0.032 0.019 0.59 2.61 The outputs from the reduction/hydrochloric acid leach process include gaseous emissions, liquid waste streams and solid waste streams. Liquid and solid wastes are, in some cases, combined in a single waste stream as a slurry or slime. The potentially toxic elements are distributed in varying amounts in certain of these streams. The waste streams are listed in Table 8.7 with the total flow, in tons/ day, representing the sum of the liquid, solid and toxic element components of that stream. Streams 67, 127 and 148 will be incinerated with other combustibles, with any residue going to the tailings pond. Stream 240 will contain the material which must be removed from the off-gases to permit discharge of the atmosphere, and may contain significant amounts of hazardous substances . Streams 30, 74 and 173 will contain one or more of the potentially toxic elements. While it is possible that other process waste streams will also contain hazardous substances, such as water treatment chemicals (streams 285, and 290), or combustion ash constituents (streams 240 and 296), these have not been considered here for reasons explained in Section 4. 8-36 en CO (U O o M Ph X o w cfl e QJ tt) i-J 0) ^-1 n3 iJ •H CO O < 0) 4-i o CO •H CO U 5: o r-l M-l -C o o o >> 5-1 Vj "O CO >^ s: 3 c CO o •H +J o 3 TS 0) Pi o H rH CO cn o & CO •H Q Cfl 4-) C 01 e -l 0) 4-1 JO S 14-1 o (1) p. >! H S _ O >, .H CO (^ TJ iH U) cfl c 4-1 o O 4-J H cn 0) w CO CO t3 U-l O c ■H tsO •H )-l O e cti cu J-l • u o CO s T3 •a C C c C O O o O Ph •H PL, •H 4-1 4-1 CO CO tfl CO M J-l &0 U C CL) c QJ •H c ■H C .H •H i-H •H •H O •H CJ CO c CO C H M H M O CN 4J u c tn C o CO < CO - ^ P tu d cu > Ph cr ,H T3 W .H 43 CJ J3 CO rH rH •H rs rH -H •"1 ^. M tn J-l rH tn CO tn u CO tn CO -O e o *> XI e o m CO CO p- > Ph en ex CO •N tn < r> XI > ej> ^ ry CO J-l hJ u »• " Xi CO > Pm pq en o tJ o tn •H ^ •H ^ ^ S rH c cu C C (U eu 3 •H CO Ai CO o ^ en ^ en CO (50 to bO ^ CO 4.) CO P. 4-1 M o Vi J-I CJ ,H o >^ O " O CO ~ CO " 00 x) O en 1 eu :: tu tu :: cu CU eu >> >, cu XI X XI bO &0 -3 60 4-1 60 XI 00 u >-l 00 •H o 3 XI XI 3 XI C XI eu XI u J-I XI 3 CO J-i 3 3 J-i 3 tu 3 en 3 3 3 3 cr x: CU CLJ rH rH tJ rH a .H 3 rH rH rH tH •H en hJ z en CO z tn en en pL, cn CO cn CO iJ < o O O O o o c^^ 00 VI (U s J-I C eu 00 J-I u tu u •H tn en c 3 eu x> (U a ex 3 •H rH PL, > J3 00 4-1 a e o C CO o 3 C CO •H 3 p:: c > eu o V4 •H ;3 J-I cn en ■H O en (U O c C en 4-) ^ cx 3 B 3 Pi (U cn eu X) O ^ CO X c e O eu O 00 4-1 tu •H eu M CO 3 M J-I K eu ^ tn rH U-i eu 4-1 rH en hJ H cn 4-1 en 3 to •H o fcu CO •H 60 o o v; tu Ph a o CO to O C 1-1 J-I X (U -H C C PQ J-I •H CQ •H CU eu M <-i CO to oo 'kof-T.l loss between See ^^nrite-up Average 100°C 100 lbs A lbs H2O surface H eat Duty a) 35 lbs H2O removed per 60 lbs solids b) Feed temp. 10°C (50°F) c) Bed exit temp. 150°C (302°F) d) Exit gas dev; point 160°F or 71°C; loss of temp from bed at 300°F to stack at 250°F e) Solids specific heat same as nickel ores = .2 BTU/lb-°F Bed Depth = 183 cm (72 in.) Superficial Gas Velocity 3.05 m/sec (10 fps) at 150°C Estimated from experience Approx. mean temp, of coastal sites Assumed Assumed Estimated • Estimated 8-A3 Plant Section: Ore Preparation and Drying Equipment Item: Function: Contingent Operation: Description: Secondary Cage Mill (two trains) - Air Swept Produce ground roaster feed, minus 0.208 mm (65 mesh) Fluid bed dryer; air cleaning system (2) 4-row cage mills per train, 60 in. , nominally 150 tph each feed constant from fluid bed dryer. Criteria Source Confidence Remarks ! Feed Rate and Size From fluid bed dryer sheet Density From primary cage mill sheet Work Index 8 Kwh/ton Brooke £. Prosser (1969) and Agarwal et al. (1976) Product Size (second cyclone tails) % Finer Than V Tyler 100 210 65 94 149 100 60 125 120 20 88 170 5 63 230 2 44 325 1 20 Air Classification a) Negligible drying occurs b) 150°C (302°F) air (inlet) at minimum of 5 lbs air/lb solids c) First cut at 65 mesh, second cut at 325 mesh, -325 mesh to precipitator. Assumed line on log screen size vs. % finer than. Beck £< Messner 8-44 Recirculating load: Assume 5% + 65 mesl is removed at first cyclone. All minrs 325 mesh to precipitator is not collected. As' precipitator exit meets California standards. Plant Section: Ore Prenaration and Drvino (continued) (continued) Criteria Source Confidence Results Product Size and Dust Entrainment Attrition produces fines to total 15% of feed at minus 4 nun (5 mesh) % Finer Than Tyler Mesh 100 25.4 1" 90 22.6 7/8" 30 4.75 4 10 2.00 9 5 1.00 16 2 210 p 65 1 44 y 325 Assumed line on log screen size vs. % fines. Confirm by cal- culation that +65 mesh not entrained at 10 fps. Assume 65/325 (1% of' feed) collected . in cyclone and -325 is all piped to oreci- pator . 8-45 Plant Section Equipment Iter-. Function Description Hydrochlorination Reactor Fluid Bed Reactor Chlorinate the Nodules The gas is reacted with nodules to form soluble chlorides. Criteria Source Confidence Remarks Extent of Reaction: Mn: 94% Fe: 27% Ni: 100% Cu: 96% Co: 100% Other Metals Assume alkali and alkaline-earth metals react at 90%: Na, K, Mg, Ca , Ba. Al does not react. Si does not react. Mo reacts at 96%. Temperature of reaction: 500°C Amount of excess HCl 100% excess Temperature of HCl inlet: 175°C Reactor configuration: fluidized bed with cyclone dust recovery and waste heat boiler. Pressure of reactor: 1 atm. Cardwell 1976a From analogy and thermodynamic data. Cardwell 1976a Cardwell 1976a and Chemical Fquilibra- tion Calculations (by energy balance) Cardwell 1976a and Industrial Practice Cardwell 1976a 8-46 Good Poor Good Good Good Process conditions specified in basic patent. No data in patent literature. Enough to subtain the reaction. Plant Section: Hydrochlorination Equipment Item: Hydrolysis/Quenching Reactor Function: Hydrolysis of FeCl3 and Quenching of Products Description: Reaction of FeCl3 to form insoluble Fe203 and to quench the nodules from 500°C to 200°C before leaching. Criteria Source Confidence Remarks Temperature of reactor = 200°C Extent of reaction of Fe complete and other metals non - (except Al which may hydrolyze in addition). Type of reactor - fluid bed with over- head recovery. Amount of air: enough to maintain low dewpoint of water at '^'75°C. Amount of cooling water: enough to furnish the reaction and quench to 200°C by energy balance. Pressure = 1 atm. Cardwell 1976a I Industrial practice Assumed Cardwell 1976a Cardwell 1976a Good Process conditions specified in basic patent. Good Good Dewpoint of gases W5°C. Good Quench to 200°C. Good 8-47 Plant Section: Equipnient Item: Function : Description: Leaching/Wash Circuit Leach Tank/Thickness Feed with Flocculant Was h the Dissolvable Solids from other Solids forming Pregnant Liquor and Tailings. Wash water is passed counter current with the solids to wash ou t valuable metals. Criteria Source Amount of aqueous HCl added to leach tank. As required to maintain pH 2 in pregnant liquor. Amount of cooling water used to cool leaching operation. As required to lower temperature to 40°C (from 200°C) and also remove heats of solution of the chlorides going into solution. Amount of recycle pregnant liquor added to leach tank - as required to make nodule slurry pumpable. Concentration of metal values in pregnant liquor. Mn 200 g/1 Ni 9.5 g/1 Cu 7.7 g/1 Co 1.8 g/1 Underflow density from thickener; 25% solids. Wash ratio (lb wash water to lb solid) : 2 to 1. Wash efficiency: = 1 fe)" Cardwell 1976a Basic thermo and standard wash prac- tice. Assumed . Cardwell 1976a Confidence Good Good Fair Good Assumed Assumed Poor where D = lbs Soln/lbs Solid in U flow F = lbs Soln/lbs Solid in feed F=9 Ds3n=6 Wash eff. = 0.986 Theoretical calcula- tion. 8-48 Fair Remarks Above pH 2, Cu will begin to precipitate. Estimated to be 20% solids. Basic patent report on pilot plant work. Fixes liquid flow rate. Solids may be slimy, consisting largely of clay-like materials. As required to give desired recovery. Also fixes metal con- centrations. Assumes perfect mixing and separations. Use 98% wash recovery. Equipment Item: (continued) Leach Tank/Thickness Feed with Flocculant Criteria Source Confidence Results Amount of flocculant: 0.1 lb/Ton of nodules and 1% water solution. Assumed. Poor Effectiveness unknown. 8-49 Plant Section: Equipment Iten: Function: Description: Cu Extraction and Wash Stripping Mixer /Settler , Surges, Organic Stripping Equipment. Remove Cu from Pregnant Liquor Copper is exchanged from the pregnant liquor with LIX65N solvent. H"*" ion is passed to the pregnant liquor. NaOH is added to main- tain pH. Organic is washed with water to assure low CI"! and organic is stripped in the normal way. Criteria Source Confidence Remarks Extent of filtering needed prior to extraction (to assure contact at 0/A interface and good separation) : to < 10 ppm solids. Temperature of system = 40°C Selection of LIX: LIX65N Extraction of Cu 99.5% Extraction of other metals = none. Standard practice Standard practice Cardwell 1976a Reported in manu- facturers literature Fair Good Fair 0/A = 2/1 pH = 2 NaOH flowrate: to maintain pH 2 by mole balance Cu extracted. Chloride scrub scheme Amount of wash water is 10% of organic flow (our guess) . Chloride purge: 1% of flow General Mills Cardwell 1976a Material balance. General Mills Demarthe et al. Extractive metallurgy Vol. II, 1976, Las Vegas AIME Meeting Assumed Good Good Fair 8-50 Poor Small amount of solids will come from wash circuit. Higher temp, causes organic evaporation. Lower temp, requires more cooling water. Some work reported on chloride in litera- ture. Unknowns are extrac- tion of other species. Life of organic in high chloride environ- ment. Chloride pro- blems in electrowinn- ing. Fixed by LIX loading available. Pilot plant results Longer purges easily obtained. Equipment Item: Mixer/Settler, Surges, Organic Strippin g Equipment (continued) Criteria Source Confidence Results Make up water to chloride wash circuit = purge flow. Stripping ratio of electrowinning: solution O/A = 4/1 achieves proper electrowinning concentration. Organic make up. 50 ppm oirganic dissolved in aqueous solutions and crud formation is based on 10 ppm of solids coining through filter and equal amount of organic entrained. Purge to organic stripping '\> 1% of total flow. H2S and NH3 flow to organic stripping: assume stoichiometric H2S, 10% excess NH3 Reaction needs assume 1% unstrippable material on organic. Material balance Standard practice Reported practice Good Fair Fair Assumed Assumed Assumed Poor Poor Poor No losses of ma- terial. Fixed by copper pick up in electrolyte. Needed to remove unstrippable metals, e.g., Co from organic Amount of unstrippable components unknown. 8-51 Operation : Copper Electrowinning Component : Starter Sheet Section Criteria Source Confidence Reni.arks 1. Blank material - titanium, 1/8" Industry trend Fair 2. Sheet size (area) - 1.00 m^ Industry typical Good 3. Sheet weight - 5.0 kg Industry typical Fair Without loops A. Current density - 200 A/ra^ Design feature Fair Within range of practice. 5. Current efficiency '- 92% Reasonable value Fair Variations in practice. 6. Deposition time - 23 h Industry typical Good 1-day cycle. 7. Cathode spacing, C/C - 0.11 m Industry typical Good 8. Cell voltage - 2.2 V Reasonable value Good Variations in practice. 9. Power consumption - 2.0 kV7h/kg Industry typical Good D.C. electric only. 10. Loop requirements - 1 sheet/5 Industry typical Good Exclusive of scrapped sheets- 11. Electrolyte circ. rate - 40 i/h/m' Reasonable value Fair Exclusive of flow- through. 12. Electrolyte temperature - 45°C Reasonable value Fair Variations in practice. 13. Demisting provisions - plastic sheet Cyrus-Bagdad Fair Foaming agents not used with LIX. 14. Anode material - Pb/8% Sb Industry range Fair Ir.proved anodes under development. 15. Cu^'^in/Cu^^out - 55/53 g/1 Design feature Fair Within range of practice. 16. H^SO.in/H.SCout - 160/163 g/1 LIX circuit 8-52 Fair a(H2S0 )determined b\ A(Cu^^) Operation : Copper Electrowinning Component : Commercial Section Criteria Source Confidence Re-.arks 1. Cathode size (area) - 1.00 m 2. Cathode weight (total) 68 kg 2 3. Current density - 18 A/m Current efficiency - 93% Cathode spacing, C/C - 0.11 m Cell voltage - 2.1 V 8. Power consumption - 2.0 k'.'fh/kg 9. Electrolyte circ. rate - 50 1/h/m 10. Electrolyte temperature - 50°C 11. Demisting provisions - poly - propylene balls 12. Anode material - Pb/8% Sb 13. Cu^'^in/Cu^'^out - 53/40 g/1 14. H2SO^in/H2SO^out - 163/175 g/1 Industry typical Determined by cathode cycle Design feature Design feature Industry typical Reasonable value Industry typical Reasonable value Reasonable value Most common practice Industry range Design feature LIX circuit Fair Fair Fair Fair Good Good Good Fair Fair Good Fair Fair Fair Includes wt of starter sheet '.•.'ithin range of practice. 7-day cycle Variations in practice D.C. electric only Exclusive of flow- through Allowance for heat- up. Improved anodes under development Within range of practice A(H2S0, )determined by by A(Cu2+) 8-53 Plant Section: Equipment Item: Function: Description: Co Extraction Mixer/Settler, Surge, H2S Reactor and Filters Remove Co from Pregnant Liquor Co is extracted with Tri-isooctylamine ,(TIDA) and stripped with water, Co is precipitated from water solution with H2S and sent to Co recovery. Water solution contains Mn and is sent to Mn recovery. Criteria Source Confidence Remarks pH adjustment: Bring pH to 4 prior to extractions. Temperature: 40°C Extraction of metals with TIDA. Assumed Co 99% Mn 1% Zn 99.5% Cu 100% Ni 0% Others 0% Organic to aqueous ratio = 2/1 Strip solution: water (slight HCl if needed) Strip/organic ratio = 1/5 give 32 g/1 Co in solution. Make up organic = 1% Amount of HCl = none. Amount of H^S in reactor = stoichio- metric to precipitate CoS Extent of reaction - complete. Temperature of reaction = 80°C Pressure is 1 atm. Standard practice General Mills (Alamine 336) Brooks & Rosenbaum Bureau of Mines No. 6159 Falconbridge Matte Leach process Above references Brooks & Rosenbaum Brooks & Rosenbaum Standard practice (by analogy) Allowance Assumed Assumed 8-54 Poor Good Fair to Poor Poor Fair Fair Poor Good Poor Higher pH probably helps extraction. May not be needed, but must be done later for Ni. Cu and Ni extraction at this temp. Little commercial practice to base this operation. Extent of Mn extrac- tion critical. Extent of Mn extrac- tion critical. TIDA strips very well. Extent of stripping of other compounds unknown. Assumed to offset nor- mal soluble losses. Unremoved Co will be recovered in Mn recovery section. Kinetics are assumed better. No need for higher pressure. Plant Section Equipr.ent Item Function Description Ni Extraction/Wash/Strip Mixer Settlers with Surges, Organic Stripping Equipment Remove Ni Preferentially from Pregnant Liquor Nickel is liquid ion exchanged from the pregnant liquor with LIX64N. V& ion going into pregnant liquor from LIX would raise pH. NaOH is added. Organic washed with water to as sure low Cl"-'- concentration. Organic is stripped in normal way. Criteria Source Confidence Remarks Choice of LIX: LIX64N Extent of Extraction Ni 99.5% Cu 99.5% Co 99.5% Others None Temperature: 40°C 0/A =2/1 Amount of NaOH addition: to main- tain pH 4 by mole balance Chloride scrub conditions: amount of wash water is 10% of organic flow. Chloride wash water purge is 1% of flow. Stripping ratio = 6.45. Organic make up: 50 ppm organic dissolved in aqueous solutions and crud formation of 10 ppm solids in pregnant liquid plus equal amount on wt. basis of organic lost. General Mills Current industrial practice. Assumed. Standard practice General Mills Estimate Estimate Calculated based on assumed loading. Estimated from reported practice. Good 8-55 Good Fair Good Fair Poor Poor Good Fair Once Co and Cu removed and pH is adjusted to 4, Ni extraction is good. Assumed to load organic to 5.5 g/1 Ni. Lower pH decreases Ni extraction. Little industrial practice. Ratio fixed by re- quirements of elec- trolyte pick up. Equipment Item: (continued) Mixer Settlers with Surges, Organic Stripping Equipment Criteria Source Confidence Results Purge to organic stripping = 1% organic flow Amount of H2S, NH3 solution to organic stripping: slight excess to remove 1% of build up of unstrippable Co from organic.' Assumed Analogy to known operations. General Mills published data. Poor Poor Removes build up of any Co which does not strip by normal means. 8-56 Operation : Nickel Electrowinning Component : Starter Sheet Section Criteria Source * Confidence Reniarks 1. Blank material - stainless steel Good 2. Sheet size (area) - 0.85 m2 Fair 3. Sheet weight - 7.5 kg Fair Without loops 4. Current density - 180 A/m^ Design feature Fair 5. Current efficiency - 96% Fair Sensitive to operat- ing conditions. 6. Deposition time - 46 h Fair 2-day cycle. 7. Cathode spacing, C/C - 0.165 m Fair 8. Cell voltage - 3.4 V Fair Determined by CD. spacing. 9. Power consumption - 3.75 kWh/kg Derived quantity Fair D.C. electrical only 10. Loop requirements - 1 sheet/5 Fair Exclusive of scrapped sheets 11. Electrolyte flowthrough - 31 1/h Fair Per cathode bug. 12. Electrolyte temperature - 60°C Fair 13. Demisting provisions - anode hoodj Fair Fair 14. Anode material - rolled lead Chemical lead deforms 15. Ni^'^in/Ni^-^out - 75/65 g/1 Design feature Fair 16. H2S0.in - pH 3.5 Design feature Good Approx. optimum pH 17. H2S0.out - 17 g/1 Process dependent Fair Determined byA(Ni2+) ,18. Na^SO, concentration - 100 g/1 Design feature Fair Required for con- ductivity. 19. H„BO^ concentration - 15 g/1 Design feature Fair Required for pH con- trol in catholyte. *Exccpt as noted, Outokumpu Oy pr actice as of 1964. 8-57 Operation : Nickel Electrowinning Component : Commercial Section Criteria Source Confidence Remarks 1. 2 Cathode size (area) - 0.86 m Fair~ 2. Cathode weight (total) - 61 kg Fair Includes wt of starter sheet 3. 2 Current density - 180 A/m Design feature Fair 4. Current efficiency - 93% Judgement Fair Sensitive to operat- ing conditions. 5. Deposition time - 164 h Fair 7-day cycle. 6. Cathode spacing, C/C - 0.165 m Fair 7. Cell voltage - 3.4 V Fair Determined by CD. spacing 8. Power consumption - 3.86 kWh/kg Derived quantity Fair D.C. electrical only 9. Electrolyte flowthrough - 11 1/h Fair Per cathode bug 10. Electrolyte temperature - 62°C Fair 11. Demisting provisions - anode hoods Fair 12. Anode material - rolled lead Fair Rolled lead deforms 13, Nn^'^in/Ni^'^out - 75/46 g/1 Design feature Fair 2+ Gives net A(Ni ) = 25 g/1. 14. 15. H sn in - pH 3.5 2 4 H^RO^out - 52 g/1 Design feature Process dependent Good Fair Approx. optimum pH. 2+ Determined byA(Ni ) 16. Na-,SO, concentration - 100 g/1 Design feature Fair Required for con- ductivity 17. H^BO„ concentration - 15 g/1 Design feature Fair Required for pH con- trol in catholyte. *Ex cept as noted, Outokumpu Oy • pr actice as of 1964. 8-58 Plant Section: EquipcienC Item: Function: Description: Manganese Recovery Heavy Metal Precipitation Remove Last Trace of Valuable Metals and Delivery to Co Recovery. Ni raffinate is heated by stream, delivered to reaction where H2S gas is bubbled in precipitating any heavy metal (Ni, Cu, Co, Zn) present as sulfides. Criteria Source Confidence Remarks Amount of H2S =1.8 MSCFM Amount of stream (sufficient to raise temperature 10°C) Liquid/solid separation. Clarifica- tion followed by centrifugation to increase solids density. Estimated Energy balance Based on current practice Fair Fair Poor Assumes efficient LIX operation To improve kinetics of reaction. Flocculant improves separation, but pre- cipitate may be diffi- cult to settle from such dilute suspen- sions. 8-59 Plant Section: Equipnent It en: Function : Description: Manganese Recovery Evaporator/Crystallizer and Drye Remove All Water from Chloride Salts in Precipitation for Elect rolysis- Metal chloride solution is evaporated to precipitate chlorides. Metal chloride solids are then sent to a dryer to r emove water of hydration as well as water adhereing to the solids in preparati on for electrolysis. Criteria Source Confidence Remarks Evaporation/crystallization: in triple effect evaporator. Drying: direct fired cocurrent rotary drier with dust removal. Amount of inerts = 100 SCFM Amount of water left on dried salts. Assume = 0.1% H2O Assumed Assumed Estimate Assumed Good Fair Poor Poor Remove sufficient water to precipi- tate chloride salts. Includes removal of water of hydration. Enough to blanket salts so that they will not readsorb any water. Significantly effects electrolysis power requirements . 8-60 Plant Section: Equipment Item: Function: Description: Manganese Recovery Fused Salt Electrolysis Electrolytic Decomposition of MnCl2 to Mo Metal and CI2 Gas. Chloride salt mixture is fed to fused salt electrolysis where electric current heats the mixture to 1200°C at base (Mn melting point) and '\^650°C at the surface and electrolytically decomposes MnClT. Criteria Source Confidence Remarks Amount of Mn production: 90% of input with balance salt wastes. The amount of CI2 production stoichiometric with Mn Gas for fugitive , control Design of reaction: Temperature Metal layer: 1300°C Salt layer: 800°C Amount/type of additive = 200 Ib/D/ Sodium tetraborate Current density = 500 amps/sq ft Potential = 6.4 to 5.0 volts energy usage is equal, 300 amp-hr 6.4/10 gm mole MnCl2 Use of electrodes = 1,200 Ib/D Cooling required of furnace = 1100 gpm of cooling water. .Amount of reverts = 5% of production Molten salt production: skimmed from furnace and cast to blocks Assumed Material balance Assumed Barton 1974 Barton 1974 Barton 1974 Barton 1974 Barton 1974 Assumed Estimated by analogy to other industrial experience. Estimate By material balance 8-61 Poor Poor Fair Fair Good Good Fair Poor Poor Poor Fair Must have same activity of Mn in salts. Need to know details of furnace design to quantify. Design of furnaces assumed similar to Bayer electrolysis. No commercial opera- tion tests. Basic patent. Basic patent Losses in commercial sized equipment un- known. Extent of consumption and fate of degrada- tion products unknown. Lining of reactor wall must be cooled. Assumes no volatili- zation at (assumed) low salt bath tem- perature. Equipment Item: (continued) Fused Salt Electrolysis Criteria Source Confidence Results Quench water flowrate to offgas suffi- cient to cool CI2 from 650°C to 125°C. By energy balance Poor Gases would be cooled in HCl recovery any- way. 8-62 Plant Section: Equipment Item: Function: Description: Cobalt Recovery Clar if ier/ Centrifuges/Leach Tanks/Evapora tor-Cry s tall izers Sulfide Precipitation Conversion of the sulfided metals to powders of Ni, Cu, Co and Zn. Criteria Source Confidence Remarks Leaching Specification: a) Slurry at 40% solids b) 70% sulfuric acid solution c) 1% purge of leach solution d) Air at 20% utilization^_ 100% conversion to SO, 4 React Stoichiometric H S with purge from leaching Nickel Reduction at 175°C and 500 psig Flash to 100°C Nickel Recovery: 90% Evaporation/Crystallizer - remove 70% of water Use 25% of NH3 solution at 100% excess prior to oxidation Cobalt oxidation at 100°C and 150 psig Cobalt Reduction at 175°C and 175 psig Assumed Assumed Assumed Assumed Assumed Assumed Assumed Assumed Assumed 8-63 Fair Fair Fair Fair Fair Fair Fair Fair Fair i Plant Section: Equipment Item: Function : Description: HCl Recovery HCl Absorption Towers, CI2 Drying Tower, HCl Stripper, H2O Stripper and Surge Tanks Recover HCl and CI2 and return Water to Process. HCl gas stream are absorbed in H2O stream, CI2 stream is "dried" using H2SO4 and CI2 product is ready for returns. HCl is removed from HCI/H2SO4/H2O mixture in HCl stripper. H2O is removed from H SO . H^O stripper under vacuum. Criteria Source Confidence Remarks First HCl Absorption Tower Flow of H2O to top of first HCl absorber. Sufficient to insure adequate driving force at the bottom of the tower, i.e., vapor pressure of HCl over HCl solution less than partial pressure of incoming HCl vapor. Amount of HCl overhead with CI2 negligible. Amount of H2O overhead with CI2: saturation at vapor pressure at 40°C. Cooling water flowrate = sufficient to remove heat of solution and to maintain temperature at 40°C. Second absorption tower same criterion as first tower only flowrate and com position are changed. CI2 Dryer Flow rate of H2SO4 to CI2 dryer = required to keep vapor pressure of H2O in H2O/H2SO4 bottom flow will be less than partial pressure of H2O in gas flow introduced in the bottom . Cooling water flowrate in CI2 dryer is sufficient to remove heat of solu- tion of H2O absorbed into H2S0^ solu- tion. Calculated Assumed Calculated By energy balance Good Calculated Fair Good Good Good Good Calculated 8-64 Good Based on HCI/H2O VLE data. Scrubbing can be very efficient. Exhaust air sent to stack scrubber for further removal of trace units of HCl, CI2. Based on H2SO4/H2O VLE data. Equipment Item: HCl Absorption Towers, CI7 Drying Tower. (continued) Criteria HCl Stripper Overhead composition (HCI/H2O) is in equilibrium with HCI/H2O/H2SO4 stream coming in at top of the tower - water content assumed to be zero. Steam delivered to HCl stripper - sufficient to vaporize the HCl. Lean HCl Scrubber Source Only small amount of water needed since HCl concentrations are low. 10 gpm of H2O allowed similar amount of cooling needed. H2O Stripper Remove water by evaporation at 0.2 bar, 175°C. Steam added - sufficiently vaporize water and allow for heat of solution change. Cooling water flow sufficient to con- dense water overhead to 40°C and cool bottoms to 40°C. Solids to waste disposal = 10 grains/ SCF of solids in all input gas streams. Make up H2SO4 required to replace acid entrained with solids purged. Calculated Calculated Estimated Confidence Results Current practice in analogous operations Calculated Calculated Fair Fair Poor Estimated Good Good Good Poor Estimated 8-65 Poor Based on HCI/H2O/ H2SO4 VLE data. Heat of solution estimated. Flows and gas com- position to tower are estimated. Based on H2SO4/H2O VLE data. Based on H2SO4/H2O VLE data. Assumptions of single purge point requires solids do not foul tower internals or heat exchanges. Losses to recovered HCl and CI2 and stripped H2O assumed negligible. Plant Section: Equipnant Item: . Function: Description: Waste Recovery Mixer Followed by Lime Boil .Tov;er and Clarifier with Filter, Absorption Columns. Recover NH3 from (NH4)2S04 by Reacting with CaO to Release NH3 - Reaction according to (NH4)2S04 + Ca(0H)2 -» CaS04'2H20-l' + 2NH3f is "released" from (NH4)2S04 with gypsum precipitated. The-NH3 is recovered in scrubber tower and gypsum is mixed wit h sludges from stack gas recovery and water is separated in clarif ier. Criteria Flow of slake lime: sufficient to react with (NH4)2S04, plus 10% excess. Recovery of NH3 assumed to be 99%. Source Report practice Confidence Remarks Steam flow to lime boil: sufficient to Calculated heat mixture to 110°C and drive off NH3 vapors with 20% excess. Cooling water overhead flow: sufficient to cool and condense overhead to 40°C Scrubbing water flow: sufficient water to make up NH3 solution of 25% to process. Liquid/Solid Separation Assume that sludge and gypsum will separate easily from water in clari- fier and will filter out of water returned to the process. Less than 10 ppm solids returned to process water. Slake lime slurry - 30% solids Slurry from settler - 60% solids Calculated Calculated Good Good Good Good More may be required for non reactive limes. Based on NH3/H2O VLE* data. Assumed Fair *Vapor-liquid equilibrium. 8-66 Plant Section: Equiprnent Item: Function: Description: Solids Waste Surge/Process Water Surges Criteria Source Confidence Remarks Solids Waste Surge Assume 2 hr surge to collect all solid waste slurry from process which are pumped to waste containment area with present slurry density. Assumed Fair Assumes all process wastes except fused salts can be disposed af in a single contain- nent area. I 8-67 Plant Section: Equipment Item: Function: Description: Material Handling Assorted Tankage and St orage Areas Provide Surge Cap acity for Plant Operating Requirements Process materials and s upplies are received fnom various sources , held in storag e, and retrieved as required. Criteria Source Confidence Remarks Nodules storage: Nodules received as 30% slurry and held wet in con- tainment pond until slurry water decanted and returned to port. Pro- vide storage for 30 days supply, reclaim as required. Coal Storage Coal received from unit train at dumping station, conveyed to coal pile for 30 days storage, reclaimed as required. Provide for dust con- trol during coal movement. Lime, Limestone Storage Materials received from train at dumping station, conveyed to storage piles (covered storage for lime) , reclaimed as required. Provide for dust control during solid movement. Process Materials and Supplies Storage Hold 30 days supply of all process materials in appropriate type of containment: pressure vessels, tankage, covered bulk storage. Assumed Fair Depends on nodules delivery schedule. Assumed Good Assumed Assumed Good Standard practice. Standard practice. Good Standard practice. 8-68 Plant Section: Equipment Item: F'jnction: Description: Services - Water Services Assorted Equipment Provide Treated Make-Up Water. Treat make-up water for cooling, potable water supply, BFN and process water surge. Criteria Source Confidence Remarks Water Services Make-up water conditioning Lime softening to remove hardness Commercial from 250 ppm to 50 ppm . Treatment chemical to cooltng tower Commercial to clean slimes and algae etc. Treatment chemical to potable water Commercial CI2 as required. BFW treatment. Ion exchange for hardfCommercial ness removal into NaCl regeneration Blowdown from cooling towers: 10 to Assumed make up. Drift: 5% of make up. Good Good Good Good Fair Depends on water quality and tower design. 8-69 Plant Section: Equipment Item: Function: Description: Services (continued) Combustion for Drying and Main Boiler/Gas Treatment. Provide drying gases and main boiler feed to produce steam and power. Burn gases and send to dryers and burn gases to provide energy for steam production and produce "in-house" electrical generation. Criteria Source Confidence Remarks Combustion: (associated with each operation requiring hot gas) . Main boiler required to generate 810 MPPHR of 185°C (11 atm) steam and 30,000 KW power generation from back pressure turbines. Gas Treatment By limestone scrubbing to take 1% sulfur and other constituents in the coal out of the gas. Water sufficient to form slurry to make sludge pumpable. Off gas disposal: high stacking. high velocity Conventional practicf in coal combustion. Calculated from pro- cess material and energy balance. Practice is conven- tional . Power generation. Good Good Good Conventional practice. Fair 8-70 Firebox and accessorie: at each location, gas cleaning done centrally. Boiler and accessories would be of standard design. Offgas composition a function of trace and minor ele- ment composition of coal. Exact requirements depends on local requirements. 9.0 DETAILED DESCRIPTION SMELTING PPOCESS 9.1 PROCESS DESCRIPTION 9.1.1 Summary Process Description Copper, nickel, cobalt and a f erromanganese alloy, if desired, can be recovered from nodules by a smelting process. A simplified block diagram of the process is shown in Figure 9.1. The nodules are first dried by direct contact with combustion gases to remove water not chemically bound to the minerals. The manganese dioxide and ferric oxide are then reduced to manganous and ferrous oxides by contact, in the presence of coke, with a carbon monoxide-rich producer gas at high tempera- ture (725°C). The hot, reduced nodules are then charged to an electric furnace, along with coke and silica. In this step most of the copper, nickel, cobalt and iron and some of the manganese are reduced (at 1425°C) and form a molten alloy phase, which separates by gravity from the unreduced manganese slag. The alloy is transferred, hot, to converter vessels where, with additional silica, the manganese and most of the iron are re-oxidized with air, separated as a slag, and returned to the electric furnace. Gypsum and coke are then added to the alloy, producing a metal sulfide "matte" phase which contains the copper, nickel and cobalt. A second liquid/liquid separation is made in the converter, with the slag returned to the electric furnace and the matte granulated by quenching it in cold water. The electric furnace manganese slag, with recycled iron-rich slags, may be further reduced (at 1480°C) with additional coke in an electric furnace to produce a molten f erromanganese alloy, if desired, which separates by gravity from the unreduced manganese slag. The f erromanganese is cast for sale, and the waste slag is granulated for disposal. The metals are recovered from the granulated matte by dissolution into strong (5%) , hot (110°C) sulfuric acid solution in the presence of oxygen (at 10 atm. pressure). The metal-bearing solution is subjected to a series of purification steps in which it is contacted with an organic medium which selectively removes the copper and nickel from the aqueous solution. Ammonia is added to the solution to control the pH during the separations. The metal values are, in turn, selectively removed from the organic fluid and transferred to acid aqueous solutions, which accumulate copper sulfate and nickel sulfate. The metal products, cathode copper and nickel, are produced from these solu- tions by electrodeposition. Cobalt is then recovered from the aqueous ammonium sulfate solution by contacting it with hydrogen sulfide, which precipitates the insoluble sulfide of cobalt, as well as small amounts of copper, nickel and other metals not removed in previous steps. The solids are removed from the aqueous ammonium sulfate solution and contacted with air and hot (100°C) sulfuric acid to selectively redissolve the cobalt and the small amount of nickel present. The undissolved sulfides are sold as minor products, and the cobalt and nickel are 9-1 :5 A ■2 I i i t o 0. 5 O [UUl in \n ul u O a: 0- o z uJ z -I o 0. lo «0 11 z z L 1 L_i I T 4 * 5 i i i 77 rzO — ^ i ^i .T r -3 i S jl: 12 It n if.^i 9-2 recovered from solution in powder form by selective reduction with hydrogen at high pressure (34 atm. pressure) and temperature (185°C). Lime is then added to the metal-free ammonium sulfate solution, and the mixture is contacted with steam (at 120°C, 2 atm. pressure) to recover ammonia for reuse in the process. The gypsum formed in this step is combined with other process solid and liquid wastes and sent to containment. Plant services include facilities for generating the producer gas used in nodule reduction, raising the necessary steam and part of the power required for process use, supplying the make-up water and cooling required, and provid- ing for materials handling for process materials and supplies. While detailed design information on the process implications involved in the smelting of nodules has not been published, enough is known about the thermodynamics of the system to permit a process outline to be constructed. Electric furnace smelting is well developed technology, and copper and nickel are currently recovered by treatment of mattes formed during the smelting of sulfide ores. Ferromanganese of high purity is currently produced directly from high quality ores. Thus, the reductive smelting of nodules to ferroman- ganese with subsequent sulfidizing of the alloy phase to form a matte is a synthesis of technologies from different areas of extractive metallurgy. Detailed information on slag properties (particularly viscosity-composition- temperature relationships) , the efficiencies of materials (coke, gypsum) and energy consumptions, and the distribution of minor metals and impurities among dust, slag, and matte phases under smelting conditions for this system, however, is lacking. The oxidative dissolution of sulfide ores and mattes is well known, but the metal separation and purification schemes are specific to nodules and are complicated by the chemical similarity of copper, nickel and cobalt. Separation and purification of copper and nickel by selective extraction with organic compounds (liquid ion exchange reagents) is currently practiced in the extractive metallurgy of copper and nickel. However, in these cases the aqueous solutions contain primarily one metal, the others being treated as impurities, not pro- ducts. Cobalt recovery from precipitated mixtures of nickel and cobalt sul- fides derived from laterites is also currently practiced. The details of the procedures used to purify the leach solutions prior to reduction, however, would differ somewhat from those used for nodules because of the differences in amount and content of impurities. The generation of producer gases from coal (or oil) for the reduction of nodules and all other plant services represent the utilization of kno\i?n technology essentially without adaptation. The detail- ed design bases used in developing the process description used are summarized in the criteria sheets in Section 9.7. While many variations are possible, particularly with respect to equipment types, for the basic scheme proposed for smelting, little information is avail- able to suggest that substantially different approaches would be more advan- tageous. Alternative routes to the recovery of copper, nickel, and cobalt from the matte, or even the unsulfidized alloy, are technically feasible. However, the route chosen attempts to make maximum use of elements of existing technology, 9-3 and variations are not likely to have major impacts on the overall plant inputs and outputs. The production of fertilizer grade ammonium sulfate from metal- free liquors would be a realistic alternative to lime boiling for ammonia recovery, and would involve buying make-up ammonia instead of lime, with a reduction in the amount of solid wastes (gypsum) produced. 9.1.2 Detailed Process Description The nodules smelting process is a combination pyrometallurgical and hydrometallurgical treatment of nodules to recover the value metals nickel, copper and cobalt, with the option of recovering f erromanganese or a storable by-product of manganese and iron. The smelting process produces a slag from which f erromanganese is recovered and a value-metal alloy matte composed primarily of nickel, copper, cobalt and sulfur. The matte is granulated and slurried and selectively leached with sulfuric acid at elevated temperature and pressure. The leach residue and metalliferous solution are separated by a series of filtering and washing stages. After liquid/solid separation, copper and nickel are selectively extracted by liquid ion exchange, stripped from the ion exchange liquid into a weak electrolyte, and recovered as elec- trowon cathodes. Cobalt is separated from the raffinate by precipitation with hydrogen sulfide and is recovered from the sulfide precipitate by select- ive leaching and hydrogen reduction along with some nickel, copper and zinc. The essentially metal-free raffinate liquor is used to granulate waste slag. Ammonia consumed in the process is recovered by lime boil and recycled to the process for use in pH control. Detailed descriptions of each segment of the process are given in the following flowsheets. (Dwg. NC-0001-SM, Figure 9.2) Wet nodules are reclaimed from storage and fed through a primary cage mill, where they are reduced to -7/8". They then pass to a feed bin for delivery to a feed belt that supplies nodules to a direct heated, fluid-bed drier for water removal. A secondary cage mill following drying further re- duces the nodules to approximately 325 mesh for delivery by the enclosed con- veyer to reduction. Mill-attached cyclones separate and recycle oversized nodules for further milling, pass on proper mesh size nodules, and deliver dusty gas to electrostatic precipitation for cleaning. (Dwg. NC-0002-SM, Figure 9.3) The dried nodules are combined with a 4.8 wt % coke, and the mixture is fed to a fluid-bed roaster for reduction with producer gas. Cyclones after the roaster remove and return large particulates, while fine dust and hot gases pass to waste heat recovery and electrostatic precipitation, which delivers the reduced dust back to the process to recombine with the reduced products of the roaster. The reduction products are blanketed by an inert gas (such as nitrogen) and delivered by drag conveyor to the smelting furnace. 9-4 <^ -J -J® -^ Jo 2 111 ill u 1 :> ? 1 ,f^ a iS'^'^ f?^ 1 tv > ■4 v^ 1 1 1 [Ulil n kr> iCiifi -Vl g*^ m •^s ^ ^ ys) ^ o kj o 1 k 5 g I I w o M 9-5 1 -A — : r^ i i ' 1 ! : : ^ 15 - ^ at M ^ .$? S ~; € '^5 r"^ 1 « j is? -- i. 1 ' - ^^ i 1 £ I/) I O I u -J. o H 9-6 (Dwg. NC-0003-SM, Figure 9.4) Reduced nodules, comprised of MnO , FeO, a small amount of metallic Fe, the value metals Ni, Cu, and Co, and the less volatile components of nodules, are smelted with silica flux at about 1425°C. An electric furnace of conven- tional design would be used. Recoveries of Fe, Ni, Cu and Co are in the range of 70 to 95% to the alloy along with minor amounts of manganese. Iron is subsequently removed and recycled to the furnace as a molten silicate. The slag is comprised mainly of manganese, iron and silica/calcium in the proper ratio for good slag fluid- ity and for subsequent production of f erromanganese. (Dwg. NC-0004-SM, Figure 9.5) I4anganese reduction to the alloy is held at a level not exceeding 1.5%, since it must be removed to less than 0.1% prior to reacting the Ni, Cu and Co with sulfur. Removal of the Mn and some Fe is accomplished by addition of quartzite in the proper ratio to produce an eutectic mixture of the low- melting silicates. Oxidation of Fe and Mn with 95% O2 conserves heat for subsequent processing. After the Fe-Mn-Si02 ^■'-^S ^^ removed as a first step, a near stoichiometric amount of sulfur must be reacted to form Ni3S2, CU2S, and C09S3. These sulfides are stable with respect to Fe and FeO at 1200°C-1400°C . The objective is to remove as much Fe as possible without a major loss of the value metals to the exhaust gas or slag. Gypsum, reduced with coke, supplies the sulfur, and a fuel oil/oxygen burner supplies the heat for the endothermic reaction. A top-blown rotary converter (TBRC) was selected for converter operations to provide intimate slag-metal-gas contact. All of the converting could be done in the same vessel, but two are shoxim to assist visualization of the four step process. The third step is to add silica flux and blow the matte-iron alloy with air until the proper amount of Fe is removed to balance its requirements in the f erromanganese (Fe-Mn) alloy that is produced in another section of the plant. Alternatively, if Fe in the nodules were low enough in relation to Mn, all of the iron silicate slag could be recycled to smelting. In all probability, early blows would be sent directly to Fe-Mn production. As a final converting step, Fe is lowered to 5% by a series of alternate blowing/slag removal/ fluxing/blowing operations. The reaction is highly exo- thermic, requiring some care in prevention of Fe304 production. In the concept- ualized process, about 2/3 of the original Fe would be recycled and 1/3 discarded in slag. The finished matte contains about 90% of the Ni,Cu and Co. It is removed by ladle to granulation. 9-7 I 1 m V5 -J 5 I s (\i ^,^ M 1 su 1^ w Pi O § I 9-8 I I [UUl V5 I I ,>^ ^ f^ ■^ f^ ? w Pi o 9-9 (Dwg. NC-0005-SM, Figure 9.6) Ferromanganese can be produced from the smelter slag in an open air electric furnace with a reductant. Coke was chosen as reductant, but several carbonaceous materials would also be suitable. The reduction reaction requires intense energy input from the electric arc. Since the charge is primarily molten, the bath surface will be exposed in molten form. Standard Fe-Mn production is from cold ore, and a crust is present on the bath surface. Medium carbon Fe-Mn is produced, and a slag containing about 8% Mn is discarded as the final waste product from the hot metal operations. (Dwg. NC-0006-SM, Figure 9.7) The molten alloy from converting is quenched in a granulation unit. The granulated matte is rake classified, with oversize granules going to wet milling for final size reduction while classifier fines overflow to a clarifier, where they are settler-thickened. The clarifier underflow recombines with wet mill effluent in a surge tank, while overflow returns to granulation. The matte slurry is pumped to an autoclave leaching vessel operated at 150 psi and 110°C. The leach products, dissolved value metals and residue, are combined with other precipitated solids and fed to a two-stage counter current rotary drum f ilter-repulp operation. The filtered solids are pumped to waste treatment while the filtrate, which is now process pregnant liquor, is pumped to a pH- adjustment circuit preparatory to selective metal removal. (Dwg. NC-0007-SM, Figure 9.8) The pH adjustment process removes the necessary residual acid concentra- tion of the pregnant liquor through the use of calcium carbonate (limestone) . Calcium sulfate, the precipitated product of neutralized pregnant liquor overflow of the clarifier is pumped to the copper liquid ion exchange process. (Dwg. NC-0008-SM, Figure 9.9) The filtered pregnant liquor passes to a three-stage counter current aqueous-organic liquid ion exchange circuit with interstage ammonia pH adjust- ment. Nearly complete transfer of copper metal to the organic exchange liquid is accomplished along with trace exchanges of the nickel, cobalt and zinc. The separated and loaded organic stream is counter currently stripped of its copper value with depleted electrolyte from copper electrowinning. The copper electrolyte is heated/cooled on passage from/to stripping to permit operation of electrowinning at a higher temperature. I'lake-up organic is added to the electrolyte-stripped reagent to offset degradation and soluble organic losses to pregnant liquor. A continuous purge to and recycle from a trace metal stripping step is operated to prevent a build-up. The aqueous raffinate stream containing the nickel, cobalt and trace metal values is sent to a neutralizing step prior to nickel recovery. 9-10 9-11 9-12 $ ^ (\I § ' § - 1 5 ^ - IM ^ ^ 1^ - fNJ m 1 ^ 1 ^ ^ (Njl ^ IM ■Xj .s^ i 5: -^ u> ; 5 » - Vj 1 ■ ' c^ O V) fT 1 V> (Si 5 (M 9-13 9-14 (Dwg. NC-0009-SM, Figure 9.10) Cathode copper is recovered from the liquid-ion exchange strong electro- lyte using conventional technology. In the stripper section, copper is de- posited onto titanium blanks to produce starter sheets which are removed, washed, looped and returned to the commercial section as starters. Full-term cathodes produced in the commercial section are washed, unloaded, and prepared for shipment as cathode to sale. The major portion of the weak electrolyte is recycled to the liquid ion exchange section for stripping, while a small purge stream is recycled to the pH adjustment process to prevent build-up of undesirable co-stripped metals other than copper. Make-up acid, used to redissolve scrap copper for return to the commercial cells for deposition, replenishes the weak electrolyte (a strong acid solution) for return to copper stripping. Sufficient steam, wash water, and make-up water are added to the circuit to offset water which vaporizes and is carried, with evolved oxygen, from the electrowinning cells. (Dwg. NC-0010-SM, Figure 9.11) The nickel-bearing raffinate from the copper ion exchange process con- tains an excess of acid, which must be neutralized with ammonia and made slightly basic and oxidized before the nickel value can be removed without co-extraction of cobalt. In addition to neutralization, tank aeration causes precipitation of iron, manganese, magnesium and aluminum initially extracted from the matte and carried with the raffinate. Because of the slimy character of the precipitated basic metal compounds, filter-aid is added and the suspen- sion is filtered. The collected precipitated solids are sent to disposal. The pH-ad justed filtrate containing some entrained precipitate passes on to nickel liquid ion exchange. (Dwg. NC-0011-SM, Figure 9.12) The neutralized raffinate is filtered of entrained precipitate, then organic liquid ion-exchanged to separate nickel in three stages of pH-controlled counter current extraction. Nickel, other trace metals, and ammonia are exchanged. The nickel-stripped aqueous phase containing primarily cobalt is directed to cobalt separation. The organic phase must be refined of entrained aqueous, dissolved ammonia, and extracted nickel and trace metals. The entrained aqueous is removed in a physical liquid/solid separation step. The ammonia is removed in two stages of recovery. In the initial stages of partial aqueous scrubbing, the organic is two-stage counter currently washed, then steam stripped of ammonia. (Dwg. NC-0012-SM, Figure 9.13) The partially ammonia-stripped organic passes to a second stage for reaction with sulfuric acid to draw the remaining ammonia into an aqueous phase. The organic-aqueous phases are counter currently contacted and separat- ed in two stages. The aqueous phase returns to nickel extraction. The organic phase, containing nickel and trace metal impurities, enters the final nickel- 9-15 9-16 Ik ^1 t> ^ V v9 ^. ^ ^ I^ **• 4 tk- \n 111 <^ ^ ^ I I VO S § § S >4^ in w I O V I I o M 9-17 9-18 9-19 producing cathode stage. This electrolyte is heated/cooled in going to/from electrowinning operations, in which the temperature is maintained slightly higher than in extraction. The organic, having released nickel and some trace metals to the electrolyte, is replenished of losses with fresh organic. A small amount of organic is purged to a trace metals removal step to prevent impurity build-up. The "cleaned" organic purge and make-up organic are com- bined with the nickel-free organic for recycle to the nickel ion exchange extraction step. (Dwg. NC-Q013-SM, Figure 9.14) The strong electrolyte is chemically conditioned with sodium sulfate and boric acid to control its conductivity and pH. Dissolved organic, carried from the liquid ion exchange step, is absorbed in an activated carbon bed prior to any electrolyte passing to electrowinning. The bed is periodically isolated from the system and steam stripped of organic. Nickel is recovered in a manner similar to that used for copper recovery. The starter sheets are picked in sulfuric acid prior to use in the commercial cells, and nickel scrap is redis- solved in ammonia-containing raffinate and recycled to the ion exchange process. The electrolyte purge required to remove impurities from the electrowinning circuit passes to raffinate neutralization. The weak electrolyte, strongly acidic, returns to the nickel stripping circuit. The nickel is shipped in cathode form for sale. (Dwg. NC-0014-SM, Figure 9.15) Cobalt, along with unextracted copper, nickel and zinc, is recovered from the nickel liquid ion exchange raffinate by precipitation with ammonium hydro- sulfide, produced by sparging hydrogen sulfide into an excess of ammonia solution. The sulfide precipitate is separated from the raffinate in a clari- fier. The clarifier overflow is filtered of entrained precipitate and sent to ammonia recovery. The underflow is mixed with purges from copper-cobalt stripping of the liquid ion exchange reagent. The mixture is pressure leached with air to preferentially dissolve the nickel and cobalt sulfides, leaving copper and zinc sulfides undissolved in the residues. The latter are removed by filtration and sold as minor products to smelters for recovery of metal value, Following pH adjustment and reprecipitation with hydrogen sulfide for final removal of any zinc and copper solubilized in the first leach, the nickel/cobalt sulfate solution is heated and autoclaved. Nickel is reduced with gaseous hydrogen. Sufficient ammonia solution is added during reduction to neutralize the acid formed. Only a portion of the nickel is removed per pass to prevent overreduction and subsequent contamination of the nickel powder with cobalt. After densif ication through repeated recycle, the nickel powder is removed, v/ashed and passed to drying and briquetting for sale. The largely nickel-free cobalt sulfate solution passes to an evaporator/crystallizer , where the remain- ing nickel and cobalt are precipitated as the double salts with ammonium sulfate. Excess ammonium sulfate is purged, and the salts are redissolved in strong cobaltic state with air to enable cobalt to remain in solution as cobaltic ammine. The stream is subsequently acidified to remove liquid salts that are separated and recycled to the pH-adjustment step. The nickel-free solution is then heated and autoclaved for removal of cobalt by hydrogen re- duction. Sufficient ammonia is added to neutralize the acid generated. The cobalt powder is dried and briquetted for sale, with the ammonium sulfate purged to ammonia recovery. 9-20 9-21 9-22 (Dwg. NC-0015-SM, Figure 9.16) The raffinate from cobalt recovery, containing process ammonium sulfate and ammonia is counter currently contacted with slaked lime in a steam strip or boil step to recover the ammonia value of the sulfate. The gypsum preci- pitate from the lime boil enters a settler, then underflows to the waste surge tank for disposal. The overflow liquor, primarily water, is recycled to the granulation/leach/washing circuit. Steam sparged into the lime boil strips ammonia, which then passes to a three-stage ammonia condenser-absorber circuit. The aqueous absorbed ammonia solution returns to the process at required locations. The ammonia-containing vent gases of the process are passed through ammonia recovery. Scrubbed vent gas is discharged to the stack for disposal. (Dwg. NC-0016-SM, Figure 9.17) Molten waste slag from f erromanganese production and converting is granu- lated by spraying it with large quantities of process waste water and make-up water, with heat removed by evaporation. Gases from smelting and converting pass through electrostatic precipitators for dust removal. Reducing and sulfur-containing gases are directed to the main boilers for combustion and subsequent cleaning; clean gases are scrubbed directly. Process dusts are conditioned by wetting and agglomerating with lime. Sulfurous dusts are recycled to the sulfidizing converter, while other dusts are returned to smelt- ing and/or purged, as required, to control the build-up of heavy metals in the f erromanganese product. All process solid and liquid effluents are collected, neutralized and pumped to the waste disposal area. (Dwg. NC-0017-SM, Figure 9.18) Facilities are provided within the plant for receiving and reclaiming raw nodules, coal, lime and limestone, silica, gypsum, ammonia, acids and other process materials and fuel. (Dwg. NC-0018-SM, Figure 9.19) Plant services include process and cooling water supply and treatment, steam raising and power generation, stack gas treatment, and producer and dryer gas production. Make-up water is clarified and softened for distribution to the process, as required. Additional treatment is required for cooling tower water make-up, boiler feed water make-up and for supplying plant potable water. Off-gas hydrogen from cobalt recovery and off-gases from reduction and smelting steps are burned in the main boiler, along with coal, to raise the required process steam and generate a portion of the power required in the process. Following particulate removal, the flue gases are combined with other process off-gases and pass to gas treatment, where sulfur oxides and other acidic con- stituents are removed by scrubbing with limestone. The scrubbed off-gases are reheated, combined with scrubbed vents from various process steps, and stacked. Gas for nodules reduction is produced in a two-stage entrained flow gasifier in which coal is mixed with preheated air and high temperature steam for the production of a CO-rich reducing gas. The gasifier product passes directly to reduction following particulate removal and energy recovery, with sulfur removal taking place after combustion along with other boiler flue gases. Coal is also used in production of hot combustion gas for nodules drying. 9-23 I 1 1^ X < :^ O > ! < o K .1 o '-M ' ^ c tj i UJ ^ ]0 ^ li ^ 1 n 9-24 — ^^ ^

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XI r-\ > -a: u w Eh in "d- o iH .H O CO ^ in o »o H O o O tN O c o -O XI O rH (N iX) in O ro O o o o M43 W Q Pj O u < u o in a, u s H 1 Di H W S CIh O CU « O EH U U W iJ W X Pi ^ () o a ro X S Tl cn u rH C) > O «: (/) H H in ro o rH rH in m Q o iH 00 ■a- CN H o O o o o h-l o O o o O O o O O o O W • • • • • CPT) X! rH > a: u W H ff\ o o o H o O fO fN "d" rH O CM O O O cn T3 cn rH > <: u < w Eh r-^ ro rH o n O CN CO CTv CM O CN O o O O o CN O O o O XI rH W Eh z o H Eh Q g D O H IX Eh 2 U H U z H 9-39 t a bi O Pj W a N Q Ti Xi iH U w H W o ro o (M n M o o c:> OJ o o o O o o o o n O CN n o o o o o CO 1 S u w rH >=? O O r\! 00 ro - m T) XI -H > < U W H ro CN C) O CM ^j- ro C3 00 ro 'd' O ro O O O O CN NICKEL "^ CATHODES TJ > U fM « ^ o u § n E 2 T) XI iH > U W H O ^ ^ CI ^ w o iH O LO Q o (Ji CN] r\j H C3 o o o J O O O O o O O O O w • . XI H > w H un •vf o iH o in o CO o ,— \ o ro c n o O o w M 1-1 o < u I/l w ^— ^ u () H o fl. o ^ H v> CM (> H ■ w 1 U w ro o kD M 9-40 o o o o o CJ O o CM < PQ O Eh U U W Pi O I— I en CNl h^ g u H 2 s o \ H Eh H ^B . PQ Q O W U K > k X3 H W Eh E-1 ^ O 00 K ^ ag O O) ro O H u 9 O O s H Pi 2 A O O 9-41 Table 9.6 Summary Flows and Distributions of Potentially Toxic Elements Smelting Process Input Rate Distribution as Percent To Plant Reporting to: Discharge Rates to: Slag Tailings Element 10^ lbs/day Slag Tailings Products 10^ lbs/da y 10^ lbs/day Barium 22.0 100 22.0 Lanthanum 8.8 100 8.8 Vanadium 2.2 5A 1 45 1.2 Chromium 0.066 55 45 0.036 Silver 0.018 55 45 0.01 Cadmium 0.11 45 3 52 0.05 0.003 Arsenic 0.29 10 19 71 0.03 0.055 Antimony 0.17 95 5 0.16 Thallium 1.15 45 4 51 0.52 0.046 Lead 2.9 45 5 50 1.3 0.14 9-42 xj •o -a -o 13 X) -a 'a ^O T3 T3 c C c c c c a C c c C G G o o O o o O o o O o o O O O H Ph •H. Ph PM ■H 4-) Ph P-, fx. p-( FM (X, P-( P-i rH U3 (X! W 03 n3 U) 05 03 0) 0) 0) 05 03 ra 00 V-i bO OO )H 00 00 00 00 00 00 00 00 U) c 0) C C (U C c C c c c c c O ■H a •H ■rH c •H •rH •H •H •H ■H ■H •H p. rH •H rH rH ■H tH rH rH rH iH rH rH rH w •H O •H ■H CJ •rH •H •rH •H •H •H •H •H •H n3 c n) (13 C ct) n3 m CC CO n3 CO (0 n H 1— 1 H H M H H H H H EH H H " rQ 05 ^ *< rH - Ph 4-1 C/3 rn H U m CJ ca Q) •s ^ S 03 ♦V ' rH QJ < -o 03 rQ 1 1 > H 1 1 1 M T3 p^ < LO 1 W Q) ». j v\ rs j 1 j t3 T3 -(3 #s •s 1 1 03 03 1 II O ;3 O C X) 03 1 1 hJ <; 1 1 1 •H iH cO U < >< o *\ #^ v\ O-i O C 00 rH r, rH #^ cO -ra 03 H M EH > m CJ i CO H 03 B to (U Jj 03 4-J 03 w 0) CJ (U O 4J ^-1 03 Pm to S 00 C LW •H o 4-i iH >> CU V-i e CO to D CO rH li-l to o •H V-i 0) QJ & +J >. CO Eh r^ O P^ iH to |ii T3 iH 03 to C ■U O O W H ^ 00 6 cO ■H rH >% c • c CO u CO CJ o u OO J-> rCl -n >. 3 rH to O 1 CO U QJ CO U 3 r (u >> >. rH tu rH >. 03 T3 OO Q) u 4-1 M 3 05 OO C/^ U rH 3 -a B u c tH c 4-1 T3 u •H P 3 •H 3 QJ 3 CO 05 3 ^ x: 3 CO U rH rH rH CX rH J-i 3 iH 03 03 rH H = 00 CO CO CO CO CJ Q to < < CO IN vO CM v^ m r-{ rH r-- O u-| •H :s V-i to •H QJ •H O C G C 03 o 12 c 6 c 4-1 4-1 6 QJ T3 O O ^-1 C/5 O o c CO CO 0) 4-1 CD •H ■H QJ 14-1 •H V-i •H C N Pi rH MH QJ 4J 4-1 rH 05 o OO 4-1 M-l S •H •H •H 00 OO o Ph cd CO •H to a CJ O 14H rH CJ O d C CO N u o o c •H Q) 05 V-i t4H cO •H m •H •H ^-1 •H •H pq •H rC CO t3 4-1 to U C 4-1 4J u QJ G U-l ^ M tJ ■H u (^ 4-1 CO QJ rH .H QJ rH o •H c o •H to X rH QJ 3 00 e QJ QJ 4J •H •H 05 •H CO !-i QJ M O rH 3 0) U •rH s 6 CO O tu CO to 4-1 O hJ hJ CO w CJ x; o iJ CO CO s W Q CJ >: CO g Vj CO QJ QJ ^ CNJ U B (J^ iJ 3 CO ;^ rH rH CM CM CO O rH rH CM rH O Csl vO LO O vn O vO r^ r^ a> CM CO CO ! ^ to 13 #. ■"^^ x) -^ rH CO vO CN ^ .H -1-1 .H O CO w ^-^ 00 O >, d CN CO •H ^^ TJ c e c p. iH CO tn •H am cl & *^^ rH O CO O 4-1 (U l-i CO e i-j 4-1 x) CO vO 4-1 O •H OJ ro W QJ .H Vj • rH O 4-1 O (U W CO CO ^— ' 4-) m CO s to ^-*\ C •X) e '-- •H •H rH a- >, (X CO 05 O -w -a 4-1 to cn ""■^ c to CM to (U (U cu vD d e o u rH O 00 (U o CO 4-1 iH 1-1 ^ C E a> w Ph QJ •H to ro a. uh 0) . (U a M ex m Jj rH iH •H C o CO 4J rH ^ X ■H u Pi CO •• — to o 4J H H 14-1 r-l B o C/D c p. to o >. O.'O •H )-i ^— ^ ^---^ 4J Vj to CO 3 CM C U rH ON O 4-1 W 4-> c e (U to to ro o rH Qj • c •H U CM o CO 4-1 csl u H c/3 •>-' to to -d to ~-^ 0) to o u ex o c u M pL, 00 CO en CM IT) r^ o CM 1 1 vD rH in vO fH CO r-i r^ 00 , g c 3 3 B o c d g CO •H •H (-1 d •H o •H 3 42 T3 B QJ •H C 6 rH •H 4-1 CO o > e CD •H rH T3 u C C t-l <-\ 13 to 4J to to CO CO to X •H to u c .c (U eq rJ > u CO CJ < < H hJ 9-44 9.5 PLANT SUPPORT REQUIREMENTS, SMELTING PROCESS It is estimated that the total plant facility, including processing and support facilities, but excluding long-term waste containment areas, will require approximately 120 acres of land. Within the processing area, the plant layout will involve equipment densities which are typical for both the non- ferrous metals and the chemical processing industries. Separate processing buildings would be provided for the liquid ion exchange operation and for the smelting and converting sections, as well as for copper and nickel electrowinning and cobalt recovery. Nodule drying and reduction, coal gasification, slag granulation and ammonia recovery would be outside installations. A boiler house/power house and banks of cooling towers would be located adjacent to the chemical process areas, and offices, laboratories, maintenance shops, warehouses and control houses, and a change house/cafeteria and parking area would also be located close to the main processing area. The aforementioned facilities would require about 30 to 50 acres of land. In addition, a considerable area is required to provide materials storage. The nodules storage and decant ponds, the coal and coke, lime and limestone, silica and gypsum storage areas, and a plant run-off and emergency waste storage area would require at least 50 acres, including margins, of relatively flat terrain. No unusually tall structures would be required in the process areas, but the main stack height could be significant, depending on local meteorological conditions. The materials movement in and out of that plant are significant: exclud- ing raw nodules and waste products, in excess of 2300 tons/day of fuels and process materials are brought in and 825 tons/day of products are shipped out. Thus, the transportation network supporting plant operations must include high- way and rail connections capable of supporting considerable traffic. The coal and coke, lime and limestone, silica and gypsum, ammonia and hydrogen sulfide, and sulfuric acid would be delivered and copper, nickel and f erromanganese would be shipped by rail. Most other materials and supplies would be trucked, except that oxygen would be purchased over-the-f ence. Manufacturing operations within the plant will be on a three-shift, 365 day per year basis. It is estimated that about 300 people will be employed in the facility, including operating and maintenance labor, supervision, and plant general and administrative personnel. The mix of skill levels would be typical of a chemical processing plant, with about one-third being skilled tradesmen and lead operators, one-third unskilled laborers, and one-third pro- fessional/management and clerical/support personnel. 9.6 ACCIDENTAL RELEASES 9.6.1 Elements Pertinent to Hazard Analysis The hot metals portions of the smelting process includes a typical metal- lurgical assemblage of equipment containing about 1000 tons of molten material. The matte produced is granulated and leached similarly to the sulfuric acid process. However, the leach vessel of 50,000 gallons capacity is pressurized to 10 atmospheres, instead of 40 atmospheres. The processes from this point utilize equipment and technology considered standard by the industry. The operating conditions for the boiler deviates from typical power plant practice 9-45 because of combustions of low heating value off-gas from reductions. The gasifier unit requires 67 atmospheres pressure. Other standard higher temper- ature, high pressure, high speed components include turbines, centrifuges, and steam heaters. Most of the process stream following leaching is subject to corrosive action of acid, ammonia, hydrogen sulfide, and/or slurry abrasion. The process inventory includes about 8 hours of molten materials in the smelting (hot metal) phase (1000 tons) ; 200 tons of material in roasting operation and about 4.5 million gallons of fluids (18,000 tons). The fluids include 400,000 gallons of acid/copper sulfate solution, 750,000 gallons of acid/nickel sulfate solution and nearly 600,000 gallons of kerosene used as ion exchange solvent. Large volume components of the system, besides the smelting furnaces and converters and ladles include: 14 Separators Clar if ier /Thickeners Roaster Leach Vessel Wet Mill Lime Boiler 2 Wash Tanks Pregnant Liquor Tank Raffinate Tanks Surge, Purge and Wash Tanks 1 Separator 80,000 gallons 50,000 gallons 30,000 gallons 3 X 10,000 gallons 180 tons 50,000 gallons 50,000 gallons 25,000 gallons 10,000 gallons 500,000 gallons 10,000 gallons 2 X 150,000 gallons 30,000 gallons 20,000 gallons 5,000 gallons 250,000 gallons 3 X 80,000 gallons 30,000 gallons 10,000 gallons 10,000 gallons 60% kerosene Reduced nodules 10% H„SO, solution I 4 Water 2% H2SO4 solution 2% H2S0^ 1% of dr y nodule wt at discharged to Electric Furnace. Criteria Source Confidence Remarks Reduction Accomplished : 100% of Mn02 ^ MnO 100% of Fe203 -^ FeO 20% of FeO -^ Fe 100% of CuO -> Cu NiO ->- Ni CoO -V CO 20% of M02O3 ^ Mo Discharge Device and Temp : Ceramic lined drag converter 650°C (1200°F) into electric furnace. Nitrogen purge of conveyer and feed bins. Feed Rate of 3185.6 tons/day May be assumed = 3200 to simplify constituent balance to reduction kiln; actual rate would approach this if reduction off gas were 50% utilized for drying. Dust Handling 20% of feed passes to cyclones, 1% of feed passes to waste heat recovery 19% to product, 0.1% of feed passes to furnaces, of which half is re- cycled to dryer and half to main boiler. Later 0.05% of feed is "lost" to flyash. Simons, Fig. 3 (1971). See also Sridhar et al. (1976) Fair Boldt et al. (1967). For conveyer, p. 265. Assumed Fair Poor Assumed Poor 9-52 Must be done in presence of solid carbon, over 700°C, to yield Fe. Inco practice. Temp, drop is estimated for 12-hr storage in in- sulated bin. Not too important and will conform to best practice. Not too Important and will conform to best practice. Plant Section: Equipment Item: Function: Reduction Fluid Bed Roaster Reduction of the value metal oxides to metals. Description: Raw dried and ground nodules are feed with metallurgical pet coke to a fluid bed roaster for gaseous reduction Criteria Source Confidence Remarks selids Temperature = 725°C Pressure Gas Temperature = 925"'C Producer Gas Composition: (wt %) CO 18.2 C02 8.1 H2 9.4 CH4 3.6 N2 49.0 H2O 11.1 H2S 0.4 NH3 0.2 Use waste heat, if possible from reduc tion roaster, to recycle entrained dust and burn combustibles. 9-53 Plant Section: Equipment Item: Function: Description: Smelting Electric Furnace Smelting Nodules Inco Thompson smelter design, rectangular in-line submerged arc 6 - electrode 18,000 KVA capacity with hot reduced nodule feed. Dimen- sions 90 ft long X 22 ft wide x 13 ft high, with electrodes- to 50 ft above top. Water cooling on bottom, charge and tap hole, jackets, and launders. Criteria Source Confidence Remarks Recovery from nodule feed to alloy: 70% Fe; 94.7% Ni; 86.9%.Cu; 85.9% Co. Max Mn = 1.5% of alloy. Electrode carbon usage reduced nodules. 30 lbs/ton Coke usage = 133% of theoretical for reduction of balance of FeO to Fe. Assume MnO -»- Mn with electrode car- bon. Sulfur in charge distributes equally to slag, metal, and off-gas. Offgas generated volume = infiltration air/volume on std cu ft/min basis. Dust froir electrostatic precipitator and waste heat recovery = 2.5% of charge. Values in dust are mostly Fe, Mn, Zn and Pb. Either market dust or melt in final slag for disposal. Slags returned at 2500°F Sridhar et al. , 1976 Modified slightly. Inspiration copper model, Yannopoulos and Agarwarl, p. 400 (1976). Derived from Doniambo practice, Boldt, Quengau, p. 405 (1967). Copper practice as basis. Yannopoulos and Agarwall, p. 390. Assumed Assumed 9-54 Fair Good Fair Fair Fair Fair Fair Fe and Mn reduction especially sensitive to operating condi- tions. Several cross-checks. Dependent upon opera- ting practice and furnace design. Also, unsure of state of Fe from 760°C dryer. Assumed 100% FeO. Some lime in the bath, plus Mn, will hold S, as will Cu in metal. An assumption that checked well with practice on well- designed, tight-fitting hoods. Plant Section: (continued) Smelting Criteria Source Confidence Results Slag composition: 37% Si02, max 9% AI2O3, balance OK for liquidus of 1200°C. Furnace heat loss taken as 25 BTU/ft2- min on walls and roof, 40 BTU/ft^-min on bottom. Number of furnaces based on 2 hr smelting time and 3 hr plug flow-thru 12" alloy depth is 4 at described size. But energy requirements are 34,500 Kw, so 2 at 18,000 Kw OK from point of view of melting. Probably take 4@ 18,000 Kva. Assumed Yannopoulos and Agarwall, V.l, 1976, p. 386. Calculation Fair AI2O3 Solubility in Rhodonite' and Fayalite is improved by con- tained CaO. Fair Required estimate for heat loss and venti- lation requirements. 9-55 Plant Section: Equipment Item: Function: Description: Oxidizing Converter - Sta ge I TBRC Lower Mn in alloy to < 0. 1% for su If idizing. 15 ton capacity on normal tilt of 17° from horizontal. Rotates at 0-40 rpm and has feed chute and oxygen lance through tight- fitting hood. One in use , one on standby. Criteria Source Confidence Remarks Flux and O2 added to reduce Mn from 1.5% to 0.1%; revert slag to smelting by ladle. Use 02 instead of air and 98% Si02 instead of sand to conserve alloy heat for sulfidizing step. O2 utilized 100%. Waste gases at slag temp, of 2500°F less 500°F AT. Assumed 1000 SCFM per unit and 3 TPD of dust. Slag composition for liquidus of about 1200°C, near-fayalite. Full cycle of 40 min, 11.6 tons of alloy/tap. Sridhar (1976) Result of energy balance also Yannopoulos and Agarwal, V.l, 1976, p. 135. Assumed Assumed Assumed, based on Good Fair Fair Good Fair To minimize loss of S to slag in next step. To maintain slag fluidity. Is prac- ticed for low consump- tion uses or heat conservation. Usual practice. Slag relatively free of other oxides. Several references. 9-56 Plant Section: Equipment Item: Function: Contingent Operations: Description: Converting Sulfidizing Converter - TBRC Sulfide production using gypsum, coke and fuel/oxygen Oxidation converter A 25 ton TBRC is charged with alloy after the Mn is reduced to < 0.1%. Gypsum and coke are concurrently added with fuel oil and oxygen, with mixing, until a stoichiometric amount of sulfur is present in the nic^kel-copper matte. Criteria Source Confidence Remarks Only enough sulfur is retained in matte to form Ni3S2, CU2S, and COgSg. Gypsum and coke required at 100% ex- cess. All excess removed in off-gas in which process gas volume = infil- tration air volume. Net RxN is, for Ni, Cu, or Co, yCaS04-2H20 + x6 + M ->■ yCaO + XCO2 + 2H2O + MxSy. Residual fuel oil/oxygen burning at stoichiometric to supply heat = heat of reaction. CO formed in bath is partially burned in the hood with in- filtration air to maintain 2500°F in flue + offset heat losses. Slag formed by gypsum is fluxed prior to 1st blow, then discarded in slag III after 1st blow. Slag discharged @ 2500°F to purge Fe. Matte composition and rate; % N± 37.0 V Cu 28.8 Pb Co 5.1 Zn Fe 5.0 Ti S 23.0 Sb Mn 0.001 As Mo 0.8 Bl P 0.002 Ag Rate =94.1 TPD % 0.005 0.005 0.003 0.0005 0.07 0.04 0.04 0.007 Sridhar et al. 1976 Good Boldt and Queneau, 1967, pp. 393- Ratio assumed. Pro- cess developed by Boliden. See Yannopoulos and Agarwal, V.I, 1976, pp. 317-330. Assumed Fair Fair Assumed Calculated Good Fair Fair 9-57 Doniambo blast fur- nace matte produced from 163% xs sulfur in gypsum. Japanese use Na2S04 waste. Could use FeS if mixing vessl is developed. 1 Available commer- cially. 40 ton TBRC being installed in Sweden to react slag with Cu concentrate, using an oil/oxygen burner . Probably have lowest metal loss this way. ^ust break up slag with Dxygen lance prior to blowing. For all molten ma- terials except Fe-Mn products. Based upon recoveries reported by Sridhar et il. 1976, and esti- nates from similar )perational results. Plant Section- Ferromanganese Reduction Equlpnient Item: Electric Furnace Function- Reduction of iron and manganese oxide slag Description: Molten slag transferred by ladle to an open-arc 9000 kW electric al furnace is reduced by injection of fine coke with prod ucer gas through the center of pre-baked electrodes. Medium- grade ferro- manganese is produced by casting, crushing and scruT? bing. Slag is granulated and dumped. .^__ Criteria T Source Confidence Remarks Required alloy composition ASTM A99-50. Sp ecs Calculated Mn 78 -82% 78.2 Fe Bs il 14.1 C < 7.5 4.0 Si < 1.2 1.0 P < n.35 0.25 S < 0.05 0.04 Ni 0.1 Cu 0.36 Co 0.03 Bal 1.92 Target slag composition Wt % MnO 9 . 6 FeO 1.2 Si02 56.5 AI2O3 10.8 CaO MgO { 20.0 1.5 MnO-Si02 1.5 FeO-Si02 3 Al203-2Si02 1.5 R0'Si02 Approx. liquidus 1200-1250°C (MgO = 0.7 CaO, mol basis) Reductant is fine coke and producer gas, plus 10 lbs electrode/ton charge. 150% of theoretical. Furnace size assumed 9000 kW, so for 50,700 kW need 6. ASTM specs and cal- culation by assumed recoveries Fair Unsure of practice of using molten charge. No known precedent. Phase diagrams Good Assumed Assumed Fair Fair "Must' be further developed 9-58 Plant Section: Equipment Item: Function: Description: Leaching/Washing - Smelting Process Granulation and Milling of Matte Prepare Matte for Leaching Matte is continuously ladled into Tund Lsh at controll ed rate. Tundish nozzle approx 3/8 in. dia. to pa 3S matte on to carbon plate, spinning at up to 100 rpm. Criteria Source Confidence Remarks Matte flow rate - 94.1 tons/day Matte temperature at tap = 2400°F. Temperature in ladle to granulation = 2300°F, from granulation = 200°F. Heat capacity of Matte: "t-H77°F 200°F 2300°F Const. XI Solids Liquid m^s^ .509 X 16 = 8.1 X 504 = 256 CU2S .363 X 15 = 5.4 X 301 = 109 COgSg .077 X 16 = 1.2 X 504 = 39 Fe .051 X 14 = 0.7 X 484 = _25 1.0 15.4 429 or AH^ . + AH^ fusion 2300 -> 200 429 - 15 = 414 BTU/lb. Maintain discharge water Temp. <_ 150°F Ball mill duty: from nominal 1/2" to 99% - 325 mesh. Slurry density 9% solids (Use nodule work index) Rate Classifier: 5% of matte goes to - 325 mesh overflow and is 100% removed at clarifier in 9% solids underflow. Material balance Standard Good Good Assumed Sridhar et al. (1976) p. 36 Assumed Assumed 9-59 Fair Good Good Poor Fair Will minimize fuming and eva- poration. Plant Section: Leaching/Wash Circuit Equipment Item: Pressure Leach Tank/Rotary Drum Filters Function: Remove extractable metal valves from smelting matte and filter dissolvable solids from other solids forming pregnant liquor and tailings. Description: Sulfuric acid and pressure sparged oxygen leach matte of metal values. Rotary drum filters remove the small leach residue. Criteria Source Confidence Remarks Approximately nine per cent solids slurry to leaching The leaching reaction involves two possible reaction mechanisms, a sul- fate and a sulfur forming species. Leach temperature: 110°C. Leach time: '^1 hr Leach pressure: 150 psi A low acid concentration of approxi- mately 75 g/1 was used to favor the SO4 reaction completely with no sulfur formation. There is a 99% dissolution on leach extraction of Cu, Ni, Co. Fe. Distribution of impurities in the leach solution: ppm ppm ppm ppm Mn: 10 Bi: 60 Pb: 90 Sb: 5000 Mo: 80 V: 5 As: 5000 P: 930 Zn: 20 Ti: 70 A 20 g/1 excess of sulfuric acid is maintained to drive the leaching reaction, Iron concentration of 5 grams per liter was assumed in the leach solution. +2 +3 Fe /Fe distribution of 1:1 Sridhar, Jones, Warner Sridhar, Jones, Warner Sridhar, Jones, Warner Sridhar, Jones, Warner Sridhar, Jones, Warner Sridhar, Jones, Warner . Sridhar, Jones, Warner . Sridhar, Jones, Warner. Sridhar, Jones, Warner. Sridhar, Jones, Warner . Sridhar, Jones, Warner 9-60 Fair Fair Fair Fair Fair Poor Fair Poor Fair Fair Fair Paper is not entirely clear on the percent of occurrence of each reaction. This is a batch leach time may not apply to continuous operation. Calculations based on article information ioes not confirm thoroughly the results reported. [here was no way of ieterming the amount of Impurities introduced by the source of sul- fur and coal used in :his literature report. Plant Section: (continued) Leaching/Wash Circuit Criteria Source Confidence Results Leach ■Residue of 1% of matte feed weight was used for guidance. Concentration of major extracted metals was maintained close to the following values: Nl = 40 g/1 Cu = 24 g/1 Co = 5 g/1 Fe = 5 g/1 Leached matte residues handled as twenty percent solids in the washing filtering step. Wash efficiency = 98%. 60% solids content of gypsum coke. Pressure of inerts 15 psi. Water criteria based on metal solu- bility of 75 g/1 Oxygen: 99% pure. Wash Water: 3 displacements Sridhar, Jones, Warner. Sridhar, Jones, Warner . Assumed Fair Fair Assumed Assumed Assumed Assumed Assumed Assumed Poor Good Good Good Fair Fair Fair 9-61 Plant Section: Copper Extraction/Stripping Equipment Item: Mixer Settlers with Surges, Organic Stripping Equipment Function: Remove Cu preferentially from pregnant liquor ^^^^ Description: Cu is liquid ion exchanged from the p regnant liquor with LIX64N . h"*" ion going into pregnant liquor from LIX would raise pH . NH3 is added. Organic is stripped of metal value with CUSO4- H2SOA solution. Criteria Source Confidence Remarks Choice of LIX: LIX64N Extent of Extraction Cu 99.5% Ni None Co None Others None One percent entrainment in organic of gaseous components Temperature: 40°C Internal = 1:1 General Mills Assumed Assumed Standard Practice Standard Practice 0/A = External = 5:1 Amount of NH3 added to maintain pH at 2.5. Mixing time: 3 min Organic make up: 50 ppm organic dis- solved in aqueous solutions and crud formation of 10 ppm solids in pregnant liquid plus equal amount of wt basis of organic cost. pH = 2 Purge to organic stripping = 1% organic flow. No. of stages required: Extraction - 3 Stripping - 2 Solvent recovery - recover entrained organic from raffinates by air floata- tion in standard cells. Calculated based on assumed loading Estimated from reported practice Cardwell 1976a Assumed Assumed Assumed 9-64 Good Poor Poor Fair Fair Fair Poor Fair Fair Rates fixed by requirements of electrolyte pickup. Removes build up of any Co which does not strip by normal means. Solubility losses unknown . Plant Section: Copper Extraction/Stripping (continued) Criteria Source Conf idence Results Floatation conditions: Residual organic: 50 ppm Promoter requirements: none Residence time: 20 min. Copper stripping No. of stages required • 2 Internal O/A: 1 External O/A: 3.1 Mixing time, min: 3.75 2 Settling rate: .15 1/min/m Percent Stripping for: Ni 0.9 Cu 87 Co 0.2 Zn 100 Others Weak Electrolyte Ni 10 g/1 Cu 55 g/1 Zn 5 g/1 Others Temp. °C 40 Surge Holding Time Electrolyte: 1 hr Organic: 4 hr Copper/Cobalt Stripping H2S required: 200% of stoichiometric requirements for precipitation of Cu, Ni, Co, and Zn. Precipitation conditions : Temperature: 40°C Pressure: 1 atm Residence time: 4 hr Wash/neutralization: O/A = 1 with: 100 g/5, NH3 Assumed Poor \garwal et al . 1976a Assumed Lab test on organic extracting simulated leach liquor. Assumed 9-65 Good Poor Fair Operation : Copper Electrowinning Comoonent: Starter Sheet Section Criteria ' Source Confidence Remarks 1. Blank material - titanium, 1/8" Industry trend Fair 2. Sheet size (area) - 1.00 m^ Industry typical Good 3. Sheet weight - 5.0 kg Industry typical Fair Without loops 4. Current density - 200 A/m^ Des.ign feature Fair Within range of practice. 5. Current efficiency ^ 92% Reasonable value Fair Variations in practice. 6. Deposition time - 23 h Industry typical Good 1-day cycle. 7. Cathode spacing, C/C - 0.11 m Industry typical Good 8. Cell 70ltage - 2.2 V Reasonable value Good Variations in practice. 9. Power consumption - 2.0 W^ni/kg Industry typical Good D.C. electric only. 10. Loop requirements - 1 sheet/5 Industry typical Good Exclusive of scrapped sheets. 11. Electrolyte circ. rate - 40 2./h/m^ Reasonable value Fair Exclusive of flow- through. 12. Electrolyte temperature - 45°C Reasonable value Fair Variations in practice. 13. Demisting provisions - plastic sheet Cyrus-Bagdad Fair Foaming agents not used with LIX. 1^- Anode material - Pb/8% Sb Industry range Fair Improved anodes under development. 15. Cu^'^in/Cu^'^out - 55/53 g/1 Design feature Fair Within range of practice. 16. H 9(1 in/H-,SO, out - 160/163 g/1 2 4 ^ t LIX circuit Fair A(H2S0^)determined by A(Cu^^) 9-66 Operation : Copper Electrowinning Component : Commercial Section Criteria Source Confidence Remarks 1. Cathode size (area) - 1.00 m 2. Cathode weight (total) 68 kg 2 3. Current density - 180 A/m 4. Current efficiency - 93% 6. Cathode spacing, C/C - 0.11 m 7. Cell voltage - 2.1 V 8. Power consumption - 2 . k\Jh/ kg 9. 10. 11. 12. 13. 14. Electrolyte circ. rate - 50 1/h/m Electrolyte temperature - 50°C Demisting provisions - poly - propylene balls Anode material - Pb/8% Sb Cu^'^in/Cu^'^out - 53/40 g/1 H2SO^in/H2SO^out - 163/175 g/1 Industry typical Determined by cathode cyci.e Design feature Design feature Industry typical Reasonable value Industry typical Reasonable value Reasonable value Most common practice Industry range Design feature LIX circuit Fair Fair Fair Fair Good Good Good Fair Fair Good Fair Fair Fair Includes wt of starter sheet Vithin range of practice. 7-day cycle Variations in practice D.C. electric only Exclusive of flow- through Allowance for heat- up. Improved anodes under development Within range of practice A(H2S0, )determined by by A(Cu2+) 9-67 Plant Section: Equipment Item: Function: Description: Copper Raffinate Neutralization Mixing-Aerating Tanks/Polish Filter To adjust the pH for nickel extracting and to oxidize cobalt. The raffinate is aerated in mixing tanks foi^ cobalt oxidation and hydroxide precipitation of some trace metals. Ammonia is added for pH adjustment and the raffinate is cleaned of precipi- tate in a polish filter. Criteria Source Confidence Remarks No. of neutralizing vessels = 4 Residence time per vessel = 30 min Mixing requirements = 1 hp/1000 gals, slurry. Polish filter efficiency - removal to 100 ppm Filter area requirement: 1 ff^/gal @ 50 lb pressure drop Wash ratio: 3 lb/lb solution Flocculant utilization: .1 lb/ton nodules . Filter wash: solids exit at 20% solids Filter aid utilization: 1 lb/1 lb precipitated solids to polish filter Vent air: saturated with NH3, CO2, H2O (? 40°C, 1 atm. Aeration requirement: 1st stage = 20 m^/rn^ slurry 2nd stage = 10 m-^/m3 slurry Assumed Assumed Assumed Assumed Assumed Assumed Assumed Assumed Assumed Assumed 9-68 Fair Poor Poor Poor Poor Poor Poor Poor Plant Section: Equipment Item: Function: Description: Nickel Extract ion/ Scrubbing/ Stripping Mixer Settlers with Surges, Organic Stripping Equipment Remove Ni preferentially from pregnant liquor. Nickel is liquid ion exchanged from the pregnant liquor with LIX64N. The NH3 going into pregnant liquor from LIX would make the solution to basic. Sulfuric acid is added. Organic is steam stripped and acid neutralized of its ammonia content. Organic is metal stripped in normal way. Criteria Source Confidence Remarks Choice of LIX: LIX 64N Extent of Extraction Ni 99.5% Cu 1% Co 1% Others None Temperature: 40°C 0/A = 5/1 Amount of H„SO, addition: to main- tain pH 4. Stripping ratio: 1/1 Organic make up: 50 ppm organic dis- solved in aqueous solutions and crud formation of 10 ppm solids in pregnant liquid plus equal amount in wt basis of organic cost. Purge to cobalt stripping = 1% or- ganic flow. No. of stages required: Extraction: 3 Scrubbing: 5 Stripping: 6 Solvent recovery - recovery entrained organic from raffinate by air floata- tion in standard cells. Floatation conditions : Residual organic: 50 ppm Promoter requirement: none Residence time: 20 min. General Mills Assumed Good Good Estimated from reported practice. Assumed Standard Practice Assumed Assumed 9-69 Fair Poor Good Fair Poor Removes build up of any Co which does not strip by normal means. Solubility losses unknown. \ Plant Section: Nickel Extract ion/ Scrubbing/ Stripping (continued) Criteria Ammonia scrub - Primary ammonia recovery from scrub solution by steam stripping 30% of scrub solu- tion. Net removal of other con- stituents: none. Ammonia scrub - secondary Mixing time: 3 mln Number of stages: 2 9 Settling rate: .2 1/min/m Residual ammonia In organic = .01 g/1 Scrub solution composition: 200 g/1 ammonium sulfate 1 g/1 sulfuric acid No. of stripping stages requires; Internal 0/A 1=1 External 0/A 5 :1 Mixing time mln: 9 2 Settling rate: .15 1/min/m Percent stripping for: Ni 98.8 Source Standard Practice Confidence Results Good Assumed Fair Cu < Co Zn Others .004 0.3 Weak Electrolyte: H2SO4 Ni Cu Others Temp. °C 40 g/1 50 g/1 .001 g/1 40 Surge holding time: Electrolytes - 1 hr Organic - 4 hr Filters Operation: removal of solids to 10 ppm. pH control will be required to prevent Co-extraction of nickel. Agarwal et al. 1976a Assumed Good 9-70 Operation : Nickel Electrowinning Component : Starter Sheet Section Criteria Source * Confidence Remarks 1. Blank material - stainless steel Good 2. Sheet size (area) - 0.86 in2 Fair 3. Sheet weight - 7.5 kg Fair Without loops 4. Current density - 180 A/ra^ Design feature Fair 5. Current efficiency - 96% Fair Sensitive to operat ing conditions. 6. Deposition time - 46 h Fair 2-day cycle. 7. Cathode spacing, C/C - 0.165 m Fair 8. Cell voltage - 3.4 V Fair Determined by O.D spacing. 9. Power consumption - 3.75 kWh/kg Derived quantity Fair D.C. electrical only 10. Loop requirements - 1 sheet/5 Fair Exclusive of scrapped sheets 11. Electrolyte flowthrough - 31 1/h Fair Per cathode bug. 12. Electrolyte temperature - 60°C Fair 13. Demisting provisions - anode hoods Fair Fair 14. Anode material - rolled lead Chemical lead deforms 15. N.-2+in/Ni^^out - 75/65 g/1 Design feature Fair 16. H^SG, in - pH 3.5 Design feature Good Approx. optimum pH 17. H SO. out - 17 g/1 Process dependent Fair Determined byA(Ni2+) ,18. Na^SO, concentration - 100 g/1 Design feature Fair Required for con- ductivity. 19. H^BO- concentration - 15 g/1 Design feature Fair Required for pH con trol in catholyte. *Except as noted, Outokumpu Oy practice as of 1964. 9-71 ^* Operation : Nickel Electrowinning Component : Commercial Section ^ irj Criteria Source Confidence Reni.arks 1. 2 Cathode size (area) - 0.86 m Fair 2. Cathode weight (total) - 61 kg Fair Includes wt of starter sheet 3. 2 Current density - 180 A/m Design feature Fair 4. Current efficiency - 93% Judgement Fair Sensitive to operat- ing conditions. 5. Deposition time - 164 h Fair 7-day cycle. 6. Cathode spacing, C/C - 0.165 m Fair 7. Cell voltage - 3.4 V Fair Determined by CD. spacing 8. Power consumption - 3.85 kWh/kg Derived quantity Fair D.C. electrical only 9. E] ectrolyte f lowthrough - 11 1/h Fair Per cathode bug 10. Electrolyte temperature - 62°C Fair 11. Demisting provisions - anode hoods . Fair 12. Anod3 material - rolled lead Fair Rolled lead deforms 13. N-i^+in/Ni^'^out - 75/46 g/1 Design feature Fair Gives net A(Ni^"'') = 25 g/1. 14. 15. H^SO, in - pH 3.5 H^SO.out - 52 g/1 Design feature Process dependent Good Fair Approx. optimum pH. 2+ Determined byA(Ni ) 16. Na^SO, concentration - 100 g/1 Design feature Fair Required for con- ductivity 17. H^BO_ concentration - 15 g/1 Design feature Fair Required for pH con- trol in catholyte. ^Except as noted, Outokumpu Oy practice as of 1964. 9-72 Plaat Section: Equipment Icetn: Function: Description: Ammonia Recovery Mixer Followed by Li:?.e Boil Tov.-ar ar.d Clarifier with Filter, Absorption Columns. Recovery NH3 from (NH4)2304 by Reacting with CaO to Release NH3 . Reaction according to (NK4)2504 + Ca(0H)2 ^ CaS04-2H,0+ + ^NH3+ is "released" from (NH4)2S04 with gypsum precipitated. The-NH3 is recovered in scrubber toxv-er and gypsum is mixed with sludges from stack gas recovery and water is separated in clarifier. Criteria Source Confidence Remarks I Flow of slake lime: sufficient to react with (NH4)2S04, plus 10% excess. Recovery of NH3 assumed to be 99%. Steam to lime boil: sufficient to heat mixture to llCC and drive off NH3 vapors out 20% excess. Cooling water overhead flow: sufficient to cool and condense overhead to 40°C Scrubbing water flow: sufficient water to make up NH3 solution flow to 2 5% NH3 to process. Liquid/Solid Se par a t ion Assume that sludge and gypsum will separate easily from water in clari- fier and will filter out of water returned to the process. Less than 10 pp'.'i solids returned to process water. Slake lime slurry - 30% solids Slurry from settler - 60% solids *Vapor-L iquid equilibrium. Report practice Calculated Calculated Calculated Assumed 9-73 Good Good Good Good Fair More may be required ^ for non reactive- slimes. Based on NH3/H2O VLE* data. Plant Section: (continued) Ammonia Recovery (continued) Criteria Source Confidence Results Stripping Conditions : Pressure: 1.5 atm Stripping ratio: 1.2 x minimum Recovery: 99% of NH3 Number of stages: 2 Feed conditioning: none Energy recovery: by vapor Condensation, tailings stripper feed/bottoms exchange raffinate, stripper. Holding Tankage : Retention time: 2 hr Temperature: 40°C Direct Condensation : Condense vent gases from indirect condenser in packed tower -by direct contact with cooled ammonia car- bonate solution at 40°C, 1.3 atm. Vapor velocity: 2 m/sec Number of stages: 1 NH3 Absorber: Absorb and condense NH3 (and residual CO2) from process vents by direct contact into cooled ammonium carbo- nate solutions. Absorber temperature: 35°C Absorber pressure: 1.2 atm Vapor velocity: 2 m/sec No. of stages: 1 Based on Nicaro practice. Standard practice. Fair Fouling tendencies of tailings unknown. Poor Based on Nicaro practice 9-74 Good Plant Section: (continued) Ammonia Recovery (continued) Criteria Source Confidence Results Vent Scrubber : Absorb NH3 from process vents into make up scrubbing water. Absorber temperature: 30°C Absorber pressure: 1.1 atm Exit gas NH3 content No. of stages: Vapor velocity: 2 m/sec. Based on Nicaro practice . Good 9-75 Plant Section: Cobalt Rf-covf-ry Equipment Tten: CI ar i f ier /Cent r If iij^fi'-i/Lf-arl-i Tank-n/ Kvaporator-Cry ■■^tall Izerg Function: Sulfide Precipitation Description: Sulflding of any trace element of Mi, Zn, Cu, Co that have remained in the raffinate qolution with Mli^WS. Conversion of the sulfided metals to powders of Ml, Cii, Co and Zn. Criteria Source IConf idence Remarks Cone, of NB3 Solution: 12.5 wt % Produce a 30% NH4HS solution Temperature rise of 60°F necessitates cooling 10% excess of S Use steam for 10°C increase of stream to 50°C Metal Precipitation : 99.9% Zn, Cu 99 % Ni 98 % Co Temperature: 80°C Ammonium Hydrosulf ide: 20% excess Clarifler Specification : Residence time: 2 hr Clarifying rate: 20 J./min/m^ Underflow density: 5% solids Overflow entrainment: 100 ppm solid Filter Specification: removal of solid to 10 ppm Wash Ratio: 2: 1 Centrifuge Specification : a) solids density of centrifuge: 16.6% b) require 50 hp centrifuge c) operate at 2000 g's Assumed Assumed Assumed Assumed Assumed Co recovery practice in laterlte Ml ore processing. Assumed Assumed Assumed Assumed Assumed 9-76 Good Good Good Good Good Fair Fair Fair Fair Fair Poor Plant Section: Cobalt Recovery (continued) (continued) Criteria Source Confidence Results Leaching Specification: a) Slurry at 40% solids b) 70% sulfuric acid solution c) 1% purge of leach solution d) Air at 20% utilization, 100% converstion to SO4 React Stoichiometric H2S with purge from leaching. Nickel Reduction at 175°C and 500 psig Flash to 100°C Nickel Recovery: 90% Evaporation/Crystallizer - remove 70% of water Use 25% NH3 solution at 100% excess prior to oxidation Cobalt oxidation at 100°C and 150 psig Cobalt Reduction at 175°C and 500 psig Assumed Assumed Assumed Assumed Assumed Assumed Assumed Assumed Assumed 9-77 Fair Fair Fair Fair Fair Fair Fair Fair Fair Plant Section: Equipnient Item: Func tion : Contingent Operations? Description: Waste Treatment Agglomeration Reduce dust loss by pelletizing Various Disc or drum pelletizers are assumed. Lime binder at 5% and water at 7% residual moisture or 10-15% of feed. ~ Criteria Source Confidence Remarks 9-78 Plant Section: Equipment Item: Function: Description: Waste Treatment Granulation Treatment of molten slags for disposal Criteria Source Confidence Remarks Nozzle flow to granular just sufficient for evaporation of water to chill slag to 77°F. Granulation overflow to clarifier is 2% solids and 9% of solid input, 100°F Clarifier underflow 90% of slag, 40% solids, specific gravity 1.4. Over- flow clear, < 5 ppm solids. Fume hood removes 1% of slag and is 20' X 10'. Assumed Assumed Assumed Good 9-79 Plant Section: Equipment Item: Function: Contingent Operations: Description: Dust Recovery All Precipitators Dust Control Various Since no particle size distributions or resistivities are known for the various dusts, we have assumed 99% removal and IOC °C temperature drop. Criteria Source Confidence Remarks Cyclones or waste heat recovery units normally remove about 50% of inlet dust . Marketability and actual analysis of dusts is unknown, but expected to approach 20% total Zn + Pb. Expect to purge either to sale or melt in slag. Assumed Assumed Poor Various references and experience as basis. 9-80 Plant Section: Equipnant Iter?.: Function: Description: Material Handling Assorted Tankage and Stora ge Areas Provide Surge Cap acity for Plant Operat ing Requirements Process materials and supp lies are received from various sources, held in storage, and retrieved as required. Criteria Source Confidence Remarks Nodules storage: Nodules received as 30% slurry and held wet in con- tainment pond until slurry water decanted and returned to port. Pro- vide storage for 3C days supply, reclaim as required. Coal Storage Coal received from unit train at dumping station, conveyed to coal pile for 30 days storage, reclaimed as required. Provide for dust con- trol during coal movement. Lime, Limestone Storage Materials received from train at dumping station, conveyed to storage piles (covered storage for lime), reclaimed as required. Provide for dust control during solid movement. Process Materials and Supplies Storage Hold 30 days supply of all process materials in appropriate type of containment: pressure vessels, tankage, covered bulk storage. Assumed Fair Assumed Depends on nodules delivery schedule. Good Assumed Good Assumed Good Standard practice. Standard practic Standard practice. 9-81 .ant Section: Function: 2£script:^n : Services - Water Services Assorted EquipTnent Provide Treated Make-Up Water. Treat inake-up water for cooling, potable water su pply, BFN an- process water surge. Criteria Source Water Services Make-up water conditioning Lime softening to remove hardness from 250 ppm to 50 ppm Treatment chemical to cooling tower to clean slimes and algae etc. Treatment chemical to potable water CI2 3-5 required. BFW treatment. Ion exchange for hard ness removal into NaCl regeneration. Blowdox-m from cooling towers: 10 to make up . Drift: 5% of make up. Commercial Commercial Commercial Commercial Assumed Confidence Good Good Good Good Fair Remarks Depends on water quality and tower design. 9-82 Plant Section: Equipnent Item: Function: Descr iotion: Services (continued) Combastion for Drying and Main Boiler/Gas Treatment. Provide drying gases and nain boiler' feed to produce stean and power. Burn. gases and send to dryers and burn gases to provide energy for steam production and produce "in-house" electrical generation. Criteria Source Confidence Remarks Combustion: (associated with each operation requiring hot gas) . Main boiler required to generate 810 MPPHR of 185°C (11 atm) steam and 30,000 KW power generation from back pressure turbines. Gas Treatment By limestone scrubbing to take 1% sulfur and other constituents in the coal out of the gas. Water sufficient to form slurry to make sludge pumpable. Off gas disposal: high velocity high stacking. Conventional practice in coal combustion. Calculated from pro- cess material and energy balance. Practice is conven- tional . Power generation. Good Good Conventional practice. Good Fair Firebox and access at each location, gas cleaning done centrally. Boiler and accessories would be of standard design. Offgas composition a function of trace and minor ele- ment composition of coal. Exact requirements depends on local requirements. 9-83 10.0 AT SEA-PROCESSING 10.1 RATIONALE FOR AT-SEA PROCESSING Possibly, the overall economics and problems of land-based nodules processing and waste disposal will be such that at-sea processing will become attractive. If at-sea technology is developed and implemented, it \>7ill have significant bearing on the Impacts of those processing operations, if any, which remain to be carried out in land-based plants. There are two main incentives to consider in processing or preprocessing nodules at the mine site, i.e., at sea. The first is the savings in transportation costs which will accrue if a more concentrated product could be shipped to a land plant. The second incentive is the cost of waste disposal for land-based processing plants. The costs of disposal of process wastes, the amount of which approximately equal the tonnage of nodules input to the plant, will be significant. In fact, problems associated with the disposal of the large amounts of wastes generated from treatment of raw nodules might ultimately constrain the development of the industry by dictating where processing plants can be located. 10.2 PROCESS POSSIBILITIES FOR AT-SEA TREATMENT Broadly, three classes of treatment schemes might be attempted for preprocessing or processing nodules at sea. They are: A minimum treatment to upgrade nodules by physical means. Partial treatment of nodules by chemical and physical means to produce an intermediate product whose volume is less than that of raw nodules. Complete at-sea treatment to produce finished metal products. 10.2.1 Benef iciation Conventional land-based treatment processes must assume that some pretreatment will have taken place at sea. Some fines will be generated during the mining and lifting of the nodules from the sea bottom and some bottom material will be brought to the surface. These materials will be rejected at sea by separations based on particle size differences. If the nodules mineralogy permitted carrying out a physical benef iciation operation at sea, it would possibly be economically attrac- tive. A low-grade tailings would be rejected at sea as waste and the product would be an upgraded nodule material, probably in slurry form, which would be shipped to land-based plants for further processing. Unfortunately, it has been repeatedly demonstrated that the nodules are not amenable to such physical benef iciations and at least some chemical processing would be required to liberate the valuable metals from the nodule's matrix. 10.2.2 Partial Processing Chemical treatment would constitute the first step in a scheme involving par- tial treatment to produce an intermediate product. Several schemes can be conceived 10-1 k which fall within this category. For instance, the nodules might be partially reduced to form a dispersed metallic phase containing the copper, nickel, cobalt and iron which would then be separated from the bulk of the manganese residue by magnetic separation. Alternatively, the nodules could be reduced if required, and/ or leached directly, to dissolve the value metals in concentrated leach liquors, which would be separated from leach residues and the residues would then be dis- carded at sea. The leach liquors (acidic or basic solutions containing metal values) could be shipped directly to a land-based plant or treated further at sea. 10.2.3 Metal Production Further treatment might involve reduction of the metal values from solution, either separately to produce fairly pure metals individually, or in a single step that would produce a mixture of metals which would require further refining. The sequence of steps which would involve reduction, leaching, separation of solution from residues and separation of metal values prior to reduction from solu- tion is, in fact, the sequence of operations which occurs in land-based plants and represents complete treatment of nodules at sea. 10.3 TECHNOLOGICAL CONSTRAINTS To be consistent with the study assumptions used as the basis for evaluations of other aspects of nodules processing, the analysis of the possibility of at-sea processing will be restricted to processes for which a technology base exists and for which some information is available. Specifically, only those processes already identified as being most likely to be used in land based plants will be considered. With sufficient development effort, additional options probably could be developed for at-sea processing which would significantly affect economics and increase the likelihood of this approach being utilized. 10.3.1 At-Sea Constraints The installation of a processing plant, or any element of that plant, in a sea- going vessel will subject the equipment to some motion. This is the principal differ- ence from a land-based plant. The motion of ships at sea depends upon the weather, vessel design and orienta- tion. Small waves would cause little vessel motion. However, over extended ocean distances (fetch) waves build to sizeable lengths and heights. These would always cause some vessel movements. Seas in expected mining areas would be subject to swells of several hundred feet length and significant wave heights of about 10 feet or less. Wave periods will most likely be between 7 and 20 seconds. The resulting vessel angular motion would normally include many degrees of roll, several degrees of pitch and yawing, and a few feet of heave and surge. These motions would produce accelerations of less than l.lg most of the time, even at the bow and stern of the vessel or high above the deck. However, motions could become much more in a few hours during a storm. These motions would include large roll angles (30°), severe pitching and slamming of the hull forward and would cause difficulty in steering and station keeping. Even under less extreme conditions, the processing plant might have to be shut down with little advance notice. Operations and/or equipment which would be most affected by these vessel motions are smelting, electrowinning, decanting, rake classification, leach thickeners, 10-2 stripping and scrubbing tanks, furnaces and converters, extraction tanks and possibly fluidized bed reactors. An artist's conception of an at-sea processing vessel is shown in Fig. 10.1. 10.3.2 Physical Benef iciation Constraints Benef iciation of nodules prior to transfer from the mining ship has been described as desirable in the discussion of slurry handling systems in the Volume II report. Benef iciation includes reduction of the nodules to small diameter particles, less than about 1.7 mm diameter, to facilitate pumping. The equipment needed for benef iciation at sea would be the same as is used in land based plants. These items of equipment, if located on the mining ship, could presumably pro- vide all the benef iciation necessary for shore or at-sea process plants. Only a few hundred tons of equipment, plus in-process material, would have to be added to the mining ship to beneficiate nodules as mined. This conceivably could be in the same treatment sequence as removal of sediment and too-small fines from the nodules in raising the material from the sea floor, where wet screens or cyclone processes would be likely methods to clean the nodules. Benef iciation steps would require relatively small amounts of electrical power which could be easily handled by the mining ship. Another possible benef iciation step, the drying of the ground nodules, could be accomplished on the mining ship in the same manner as for on-land process plants. More fuel for hot combustion gas drying would be needed at sea. This benef iciation step is eventually required for all processes, except the high-temperature sulfuric acid and cuprion processes which utilize wet fines as feed. None of these benef iciation treatment methods are practical for an installation on the transport vessels or barges. These vessels would have much less time to accomplish the benef iciation since the benef iciation could only be performed while the transport vessel is steaming to port and not continually as on the mining vessel. Therefore, the benef iciation would have to be performed at approximately twice the mining rate for southern California destined nodules. Also, all of the transport vessels would have to be equipped to benef iciate, which would entail much higher equipment costs. Therefore, benef iciation at sea would probably either be done on the mining ship or not at all. 10.3.3 Processing Constraints In the land-based process routes which have been evaluated, all value metals are dissolved in the pregnant liquor from leaching operations and must be selec- tively removed prior to reduction to permit the direct production of marketable metals. The use of liquid ion exchange reagents, which are selective for the appro- priate metal at the process conditions used, has been assumed with contacting and phase separation taking place in conventional mixer-settler units. The size and cost of these units is dominated by the phase separators which must be quite large to allow adequate disengagement. Since these separators depend on gravity to achieve phase separation, the operation would be adversely affected by fluid motions induced by the pitching and rolling of the processing vessel, even if the degree of motion were quite small. Therefore, units of conventional design may not perform adequately in this environment. Two alternatives to the use of conventional liquid 10-3 a. •H bO C •H C •H o 4-J 0) en o A1, c •H cn w 0) u o S-i p-l 0) CO < &0 lC-4 ion exchange mixer-settlers are available which would alleviate this problem. One alternative involves the use of coalescers, such as knit-wire mesh elements, which greatly reduce the size of the separators and might permit phase separation to take place if pitch and roll are not severe. These units have not been demonstrated under at-sea conditions however, and are not in general use in land-based plants. A second alternative involves the use of solid ion exchange resins. Since metals are loaded and extracted alternately from single fluid phases, the separation problem will not exist. However, solid ion exchange technology is not used on land-based plants because of the high costs involved and their use at sea would be subject to the same limitation. Even if adequate ion exchange separation technology could be developed, problems would still occur in the metal reduction step. On the land-based plant is is con- ventionally done by electrowinning for copper and nickel and by selective leaching of sulfide precipitates for cobalt. Cobalt recovery presents no problem in this case since the relatively small amount of sulfide precipitate could be shipped to land-based plants for further processing. However, the operation of a conventional electrowinning tank house on seagoing vessels is difficult to envision because of the problems associated with vessel motion. A modification of the conventional cell design can be envisioned to permit anodes and cathodes to be stabilized against the ship's motion. However, this would require the development of new technology, not only in hardware design, but possibly in solution treatment as well, to prevent the occlusion of impurities suspended in the electrolytes which normally fall to the cell sumps under the influence of gravity. As an alternative to complete treatment to produce metals of cathode specifica- tion, the pregnant liquors might be treated to reduce selectively or totally the value metals with the production of materials which would require further refining. For example, the strip solutions from a modified ion exchange step could be reduced chemically, with hydrogen or hydrogen sulfide for instance, to avoid problems in- volved in the redesign of the conventional tank houses. Alternatively, the entire metal content of the pregnant liquor could be reduced in a single step and the resulting precipitate redissolved in a land-based plant where metals separation and reduction would occur in a conventional manner. Such an approach, however, would involve the transport of a large mount of reductant from land to the processing facility. The use of electricity as a reductant cannot be regarded as an efficient use of energy, but is convenient from a processing point of view, since all that would be required is the transportation and combustion of fuel. Alternative reduc- tants, such as hydrogen, hydrogen sulfide, and carbon monoxide, would more likely have to be manufactured at sea from raw materials transported from land. Thus, transportations cost savings which would result from shipping highly concentrated, impure metal precipitates would be partially offset by the costs of transporting reductants. An alternative to reduction at sea is to transport the pregnant liquors pro- duced at sea to land-based plants for metal separation and reduction. However, with the exception of the reduction/hydrochloric acid leach and smelting processes, the pregnant liquors produced in the hydrometallurgical portions of the processes are very diluted, containing metal values in approximately the same concentration as the raw nodules themselves. Thus, except for the fact that the shipment of clear solutions would be somewhat easier than the shipment of nodules slurries, savings would be unlikely in transportation costs for this approach. A further complica- tion arises from the fact that large volumes of solution must be made up at the 10-5 at-sea processing plant Lo replace the liquid shipped to shore. If seawater cannot be used directly, which is unlikely in view of the potential corrosion and fouling problems, either the spent liquors will have to be returned from the land-based plant or a large desalinization operation will be required at sea. In either case, the cost impact will be significant. An at-sea smelting route offers a potentially significant advantage in both transportation and waste disposal costs in that the metallic phase produced in the first smelting step is very highly concentrated in copper, nickel and cobalt, amounting to less than 3% of the original nodules weight. The ferromanganese produced in the subsequent reduction step is also highly concentrated in the sense that is is the final product desired and at the required purity. Thus, there exists a potential for significant savings in both transportation and waste dis- posal costs for the smelting route. However, severe technological problems exist in attempting to carry out smelting operations at sea. These problems arise because the smelting operation depends upon gravity to effect phase separations. Even on land-based plants, the slag composition must be carefully controlled to achieve sufficient fluidity to permit the reduced metallic phase to separate in reasonable holding times so that it can be removed for further processing. Any metallic phase which is not removed from the slag in the first smelting step is likely to be removed in the second step and result in both a yield-loss of copper, nickel and cobalt and a contamina- tion of the ferro manganese product. Since little is known about the physical properties of slags which would be produced in manganese processing and, since significant technical and safety problems are anticipated in attempting to carry out a smelting operation at sea given the expected vessel motion, the smelting route is not a likely candidate for at-sea processing at this time and is not con- sidered further. In summary, complete at-sea processing of nodules would require the develop- ment of new technology, particularly in the areas of metal separation by ion exchange and in metal reduction by electrowinning. While at sea use of a smelting route is not likely, preprocessing at sea to produce either pregnant liquors which would be shipped to land-based plants for further processing or for the production of impure, mixed metal precipitates is possible. Both approaches would offer the potential of reducing the land-based waste treatment costs, but a careful economic analysis would be required to determine the relative savings in transportation and processing costs for each route. At-sea production of a mixed metal precipi- tate will greatly reduce sea-to-land transportation costs, but would increase land-based metal processing costs relative to a scheme which involved shipment of pregnant liquors to land plants for metals separation and selective reduction. It is not likely that shipment of relatively dilute pregnant liquors to a land plant will reduce transportation costs, but the processing sequence will be relatively simple. 10.4 SELECTION AND DESCRIPTION OF POSSIBLE AT SEA TREATMENT SCHEMES 10. A. 1 Partial Processing at Sea Based on the rationale developed in Section 10.3, those portions of the Reduction/ Ammonia Leach,. High Temperature Sulfuric Acid Leach, Reduction/ 10-6 Hydrochloric Acid Lea' h and Cuprion/Ammoniacal Leach processes flow sheets con- sidered potential]^' adaptable for use at sea were defined (The smelting process is not included). I'he modified processes are shown in Figures 10-2 through 10-5 which are, essentially, altered versions of the flow sheets for land-based plants with pregnant liquor produced at sea and processed on land. The portion of the process conducted at-sea is shown in the shaded area. The land portion is unshaded. Where a function is performed both on land and at sea it is shown by a diversion line splitting the function. The processes would be modified slightly both at sea and on land for cases in which metal mixtures are reduced from solution at sea. Equipment requirements and sizes were estimated from information contained in the detailed flow sheet and line tables and the criteria summary sheets. The list of major equipment units, number of units required and approximate size of each unit onboard the sea process plant for each possible at-sea process is shown in Tables 10.1, 10.4,10.7, and 10.9. The list of equipment differs slightly from the detailed land plant process flow sheets. The ammonia leach thickeners are replaced by centrifuges. The tailings washings are performed by banks of mechanical centrifuges with intermediate surge tanks in all processes. In the sulfuric acid and cuprion processes, the rake classifiers are replaced by two banks of cyclones and the counter-current decanting operation is replaced with trains of five centrifuges alternating with surge tanks. The equipment units noted above perform the same function as the units listed on the original flow sheets but are better suited for use at sea either because of reduced size or adaptibility for use where the operation will be subjected to motion. The lists of equipment does not include boilers, distillers, electrical generators, tankage requirements in the hull, nor pumps, piping, and vents and similar small ancillary equipment. Certain nodule benef iciation steps are assumed to be carried out on the mining ship prior to transfer to the processing vessel, specifically, removal of sediment, mud and other fine wastes and size reduction. Only a few hundred tons of equipment is required to perform these operations. Drying of process feed, i.e., ground nodules, is required for the Reduction/ Ammonia Leach and Hydrochloric Acid processes. For the purpose of this evaluation, drying is assumed to take place on the processing vessel since, in part, waste heat from other on-board operation can be utilized. It is assumed that slurried ground nodules would be pumped to the processing vessel and, after dewatering, fed directly into a dryer. 10.4.2 Reduction/Ammonia Leach Partial Process Options The at-sea Reduction/ Ammonia Leach Process for the production of pregnant liquor (see Figure 10.2) includes: ore preparation, reduction, cooling, leaching and washing of the nodules, the steam, power and producer gas generation and ammonia recovery elements of the land plant. In addition, desalinization of sea water by distillation has been provided, and coal fuel has been replaced bv fuel oil. The pregnant liquor is transported to shore for LIX extraction and stripping, nickel and copper electrowinning, and cobalt recovery. Ammonia recovery is also necessary on-shore, and the raffinate is returned to the at-sea portion of the process plant. 10-7 10-8 CD Z CO CO < LU LiJ O CO o j ■< < Q- lU It 5^ 2 R 0- ^fi h- Z < _J CL m Q o Z I— 1 R^ < _J w dJ 1— 1 ^ fi, J 50! 0) .8 10-9 i§ :3 a: J cc o •^ as z _j O C/3 CO LU o o CE CL o o X Q. ui QC CL ^ o 2? ,l5l •^■^1 o ■ — ul ? c:0 — •i .J Z I o o M 5 1- ' < , , _j V Q_ ^ I z 4 — ► < _J rz=.:.r.:=r..cQ s °2 O o P(l. 1«1 ■«a 10-10 (S> •z. (S) < CO UJ Ijj en o CJ ^ q: CL LJ cr Q. J :; a 52 2fl I- 10 \D 5 rt a 5 t UU .1 a: C) LJ v.; Oi:: ^^^^^^kl ' :r- -^ ^ 1 ^ i2 ^ig TV 55 i " o a z ^- o UJ o q: 0- u o g^ UJ DC < UJ < _J Q_ z: < I o w a; u h- 1 10-11 Table 10.1 lists the specific at-sea process plant units that would be appro- priate for the Reduction Ammonia Leach process. As noted before, the ammonia leach thickeners are replaced by centrifuges and tailings washings at sea is performed by banks of centrifuges separated by surge tanks. Table 10.1 Major Equipment Requirements and Approximate Sizes for Shipboard Partial Processing Plant Reduction/Ammoniacal Leach Process NUMBER UNIT Mill Feed Bin Primary Cage Mill Dryer Feed Bin Fluid Bed Dryer Secondary Cage Mill Precipitators Electrostatic Precipitators Dried Nodule Tank Bag Filter Fluid Bed Reduction Roaster Cyclones Fluid Bed Cooler Cyclone Waste Heat Recovery Quench Tank Electrostatic Precipitators Aerators Thickener Centrifuges Surge Tanks Tailings Wash Centrifuges Surge Tanks Tailings Stripper from CCD Wash NH Absorber " Condenser CO Absorber NH Absorber Vent Scrubber Anhydrous Ammonia Storage DIMENSIONS, OF EACH UNIT Length Width Height 20' Diameter 40' 10' 1 10' 20' t 40' 25' ' 30' 10' 1 10' 8' If 15' 20' 10' 20' 10' Diameter 20' 15' 15' 15' 25' Diameter 40' 8' It 15' 25' " 40' 10' n 20' 20' 10" Diameter 10' Diameter 20' 24' 12' 24' 20' Diameter 20' 10' It 10' 10' " 10' 20' 8' Diameter 15' 5' 11 15' Diameter 30' 15' 5' Diameter 15' Diameter 30' 12' ti 40' 30' II 40' 12' " 25' 50' 45' 20' 5' Diameter 20' 12' ti 25' 12' ti 25' 60' 12' Diameter ction 12' Diameter 40' 10' 4' Diameter 15' Diameter 30' Additional Equipment Required Pregnant Liquor Stripper BMC Centrifuges Surge Tanks 10-12 Table 10.2 summarizes the requirements for services including power, steam, cooling water and process water for the production of pregnant liquor at sea by the Reduction Ammonia Leach process. Table 10.2 Reduction/Ammoniacal Leach Partial At-Sea Processing Pregnant Liquor Production Services Profiles Power Steam Function (kW) 103 Ib/hr Ore Preparation & Drying 3,700 Reduction 4,500 Leaching 3,650 Tails Washing 3,450 NH3 Recovery 500 250 Materials Handling 100 Services 2,500 151 Cooling Process Water Water (gals/min) (gals/min ) 1,400 26,000 Totals 18,400 401 27,400 500 7 165 620 1,300 If the pregnant liquor produced in the at sea plant is steam stripped to remove ammonia, in equipment similar to that used to strip the tailings, the value metals will be precipitated as a mixed basic metal carbonate (BMC). This could be removed from the liquor by centrifuging and sent to shore for further processing while the metal free solution would be recycled on board. The additional equipment required for this treatment step is summarized in Table 10.1 and the plant service profile in Table 10.3. Table 10.3 Reduction/Ammoniacal Leach Partial At-Sea Processing Basic Metal Carbonate Production Services Profile Power Function (kW) Ore Preparation & Drying 3,700 Reduction 4,500 Leaching 3,650 Tails Washing 3,450 NH3 Recovery and BMC Precipitation 850 Materials Handling 100 Services 3,500 Cooling Steam Water 1 03 Ib/hr (gals/min) 1,400 Totals 19,800 510 151 671 53,000 54,400 Process Water (gals/min) 500 7 1,243 620 2,370 10-13 10.4.3 High Temperature Sulfuric Acid Leach Partial Process Options This process is shown in Figure 10.3 for the production of pregnant liquor and includes in the at-sea process plants the functions of wet ore preparation, sul- furic acid leaching, tailings washing and desalinization. If pregnant liquor is shipped to shore for metal removal, stripped raffinate, sulfuric acid and fuel are carried back out to the partial-process plant. The land plant performs the same metal extraction functions as described for the ammonia process system, except in a manner appropriate for an acid sulfate system. Table 10.4 lists the major units used on the vessel plant and estimates of their size. The sulfuric acid process requires a relatively small plant compared to the Reduction/Ammoniacal process since the drying and reduction steps are eliminated. Table 10.4 Major Equipment Requirements and Approximate Sizes for Shipboard Partial Processing Plant High Temperature H2SO4 Leach Process UNIT Feed Bin Cage Mill Surge Bin Rod Mill Cyclones Surge Tank Iron Refuse Scrap Preheater Heater Leaching Tank Reactors High Pressure Flash Low Pressure " Vacuum " Centrifuges Tanks Banks NUMBER UNIT DIMENSIONS EACH Length Width Height banks of each 20- Diameter 20' 20' 12' Diameter 25' Diameter 20' 25' 12- Diameter 20' 4' 10' 25' Diameter 25" 20' 20' 20' 25' Diameter 25' 5' 15' 15' 40' 10' 10' 15' 20' 20' 30' 20' 10' Diameter 20' Diameter 20' Additional Equipment Required for Metal Sulfide Production Cu,Zn Precipitation Reactor 2 Centrifuges 4 Neutralization Tank 1 Centrifuges 4 Ni, Co Precipitation Reactor 2 Centrifuges 4 Surge Tanks 2 Lime Boil Vessels 3 Lime Slaker 1 Centrifuge 4 NH3 Vent Scrubber 1 H2 Production Plant (including H2S) 14' Diameter 30' long 14' 6' Diameter 20' Diameter 25' 14' 6' Diameter 14' Diameter 30' long 14' 6' Diameter 25' Diameter 30' 20' Diameter 20' 20' 15' 25' 16' 8' Diameter 5' Diameter 20' 30' 40' 40' 10-14 Table 10.5 summarizes the services required in the at-sea portion of this process for the production of pregnant liquor. T able 10.5 High Temperature Sulfuric Acid Leach Partial At-Sea Processing Pregnant Liquor Production Services Profile Function Ore Preparation Leaching Tailings Washing Materials Handling Services Power (kW) 2,900 1,350 4,600 100 1,450 Steam 10^ Ib/hr 383 Totals 10,400 383 Cooling Process Water Water (gals/min) (gals/min) 29,600 29,600 25 775 800 Reductions of impure metals from this pregnant liquor at sea is more diffi- cult than for the ammonia based processes. Either hydrogen or hydrogen sulfide are the only practical reductants; the latter can be used at lower pH's and pressures and give high recoveries although its use on board ship presents obvious safety problems. While H2S might be purchased on land and shipped to the at-sea plant, a lower on-board inventory would be required if sulfur and propane/butane (for hydrogen manufacture) were shipped out and H2S only made up on demand. In addition, limestone and lime would have to be shipped to the at sea plant to neu- tralize the acidity formed in the precipitation reactions in order to allow the complete recovery of cobalt and nickel which process economics would require. The additional equipment required for mixed metal sulfide production is summarized in Table 10.4 and the plant service profile in Table 10.6. Table 10.6 High Temperature Sulfuric Acid Leach Partial At-Sea Processing Mixed Metal Sulfide Production Services Profile Function Ore Preparation Leaching Tailings Washing H2S Production & Metal Sulfide Precipitation NH3 Recovery Materials Handling Services Totals Cooling Process Power Steam Water Water (kW) 103 ib/hr (gals/min) (gals/min) 2,900 1,350 383 29,600 4,600 25 700 42 1,600 88 2.700 50 4,300 178 150 1,600 475 900 14,000 35,500 1,200 10 -15 10.4.4 Reduction/Hydrochloric Acid Leach Partial Process Options The at-sea portion of this process for the production of pregnant liquor is shown in Fig. 10.4 and includes the ore preparation and drying, hydrochlorination and reduction, hydrolysis and quenching, leaching, washing, HCl and chlorine recovery, and washing elements of the land plant. In addition, desalination, and chlorine burning to produce HCl is added to replace the exchange, performed on land, with a separate plant. The land plant includes metals separations and electrowinning sections plus chlorine salts drying and fused salt electrolysis to obtain manganese metal. Chlorine recovered in the land plant is liquified and shipped back to the sea plant as are fuels and either propane or butane for hydrogen production. This process does not ship large volumes of raffinate back to the at-sea plant when producing a preg- nant liquor at sea. Table 10.7 lists the units of equipment needed at sea, wherein the land plant design is altered to replace thickeners with six banks of centrifuges with intermediate surge tanks for tailings washing. Required services are summarized in Table 10,8. Table 10.7 Major Equipment Requirements and Approximate Sizes for Shipboard Partial Processing Plant Reduction/HCl Leach Process UNIT Mill Feed Bin Dryer Feed Bin Primary Cage Mill Fluid Bed Dryer Secondary Cage Mill Cyclones Electrostatic Precipitators Dried Nodule Tank Hydroclorination Reactor Hydrolysis/Quench Reactor Electrostatic Precipitators Leaching Tank Tailings Washings Surge Tanks HCl 11 Absorption- -1 -2 CI. Drying ci; Surge HCl Surge Stripping H O " ^ ti H SO Surge Aqueous HCl surge Make up H^SO^ 2 4 Lean HCl Scrubber H2 Production Plant HCl Production Plant NUMBER 1 1 1 1 1 3 1 1 1 1 1 1 6 banks of 8 centrifuges 5 1 1 1 1 1 1 2 2 1 several 1 1 UNIT DIMENSIONS EACH Length Width Height IS- Diameter IS' IS' II IS' 20' 10' Diameter IS' Diameter 30' 20' 10' D iameter 3' Diameter 10' 10' 10' 10' 20' Diameter 20' 2S' If 3S' IS' ti 30' 10' 10' 10' 20' Diameter 20' IS' Diameter 15' IS' II 2S' 10' " 20' 7' II 15' 20' II 20' 60' II 30' 12' Diameter 20' 2S' II 40' 10' II 10' 15' ■1 15' very large IS' Diameter 15' S' Diameter 10' 60' 80' 40' 30' 30' 30' 10-16 Table 10.8 Reduction/Hydrochloric Acid Leach Partial At-Sea Processing Pregnant Liquor Production Services Profile Function Services Power Steam (kW) 103 Ib/hr 2,300 1,100 1,700 16.9 1,650 660 100 1,000 160 3,750 Cooling Process Water Water (gals/min) (gals/min) Leach Wash 2,300 6,900 500 Ore Preparation Hydrochlorination 1,700 16.9 6,800 11, HCl Recovery 1,650 660 72,200 Materials Handling 100 Cl2 Burning 1,000 160 10,000 460 Totals 11,600 840 96,000 980 The pregnant liquor produced by this process is a very concentrated mixture of chloride salts of manganese, heavy metals, and alkali earth and alkali metals whose weight and volume is nearly equal to that of the raw nodules. It does not appear likely, however, that it will prove to be economically feasible to selectively reduce the value metals (manganese, copper, nickel, and cobalt) from this solution to reduce transportation costs still further. 10.4.5 Cuprion-Ammoniacal Leach Partial Process Options The prospective at-sea and land partial process plants for the Cuprion process are shown in Fig. 10-5 for the production of pregnant liquor at sea. Functions accomplished at sea include ore preparation and reduction, reduced pulp separation, leaching, washing, ammonia recovery, CO gas generation, and oxygen and fresh water plants. Pregnant liquor is sent to shore, and the stripped raffi- nate returned. The shore plant is similar to the Reduction/Ammonia Leach process except that selective raffinate stripping is required to produce the high strength ammonia solution needed for reduction. Table 10.9 lists the equipment units appropriate for partial processing at sea and Table 10.10 contains the services profile for the Cuprion process at-sea plant for the production of pregnant liquor. As is the case for the Reduction/Ammonia Leach process another option exists for at-sea processing, involving pregnant liquor stripping for the production of a basic metal carbonate cake (BMC) which would be shipped to shore for metals separa- tion and recovery. The stripped liquors would be recycled at sea, but a selective raffinate stripping operation would be done at sea to provide the required reduc- tion liquor composition. This would be combined with BMC stripping. The additional equipment required for this treatment step is summarized in Table 10.9 and the plant service profile in Table 10.11. 10-17 Table 10.9 Major Equipment Requirements and Approximate Sizes for Shipboard Partial Processing Plant Cuprion/Ammoniacal Leach Process UNIT Feed Bin Cage Mill Surge Tank Rod Mill Cyclones Surge Tank Reactors pH Controller Cooling Units Aeration Tank CCD Centrifuges Tanks NH Absorber Tailings Strippers " Coolers Vent Scrubber/CO NH Absorber Gasification Acid Gas Removal H Removal Oxygen Plant CO Plant O Compressor LOX Tanks NUMBER 1 1 1 1 2 1 6 1 7 3 15 6 1 1 1 1 1 1 2 2 1 1 2 2 banks UNIT DIMENSIONS EACH Length Width Height 20' Diameter 20" 20' 12' Diameter 25' Diameter 20' 25' 12' Diameter 20' 4" 10' 25' Diameter 25* 50' II 50' so- II 30' lo ' 10" 10' 15' Diameter 15' 20' 10' Diameter 20' Diameter 25' 10' II 20' 20' II 50' 20' 17' 17' 10' Diameter 20' 10' II 20' 20* II 30' 10' " 30' 10' II 30' 60' 60' 40' 60' 85' 40' 30' 18' 20' 12' Diameter 60' Additional Equipment Required for BMC Production Pregnant Liquor Pre-Stripper Pregnant Liquor Final Stripper BMC Centrifuges Surge Tanks 14' Diameter 40' 18' Diameter 40' 10' 4' Diameter 25' Diameter 30' 10-18 Table 10.10 Cuprion/Ammoniacal Leach Partial At— Sea Processing Pregnant Liquor Production Function Ore Preparation Reduction/Leaching Leaching Washing NH3 Recovery Materials Handling Services Oxygen Plant Services Profile Power Steam (kW) 10^ Ib/hr 3,200 4,000 700 4,450 500 350 100 6,450 100 9,600 Totals 29,000 450 Cooling Water (gals/min) 43,000 31,000 4,000 78,000 Process Water (gals/min) 900 908 Table 10.11 Cuprion/Ammoniacal Leach Partial At— Sea Processing Basic Metal Carbonate Production Services Profile Function Ore Preparation Reduction /Leaching Leaching Washing NH3 Recovery and BMC Production Materials Handling Services Oxygen Plant Cooling Process Power Steam Water Water (kW) 103 ib/hr (gals/min) (gals/min) 3,200 4,000 43,000 700 4,450 8 850 1,100 97,000 1,360 100 13,750 100 6,000 900 9,600 Totals 36,700 1,200 146,000 2,270 10-19 10.5 PROCESSING VESSELS The ship or barges on which a processing plant is installed will probably operate closely connected to the mining ship. The processing vessel would need at least annual trips to port for repairs and refurbishment. This would probably be accomplished at the same time as the mining ship. With adequate power and stowage capacity for supplies, the processing vessel may operate while underway between the mining site and port. 10.5.1 General Physical Requirements The at-sea processing vessels need both space and weight capacity for each processing plant to accommodate Processing plant equipment. Nodule storage, - Reagents and supplies storage, Steam generation. Power plants for electrical generation. Fresh water plants, Hotel accommodations for the crew. Fuel and equipment, Transfer equipment to transportation vessels. 10.5.1.1 Processing Plant Equipment At the maximum, any vessel processing plant would need adequate space for the installation of the major equipment items such as described in Section 10.4, the auxiliary facilities such as pumps and piping, and the utility services such as boilers, distillers and power plants. For safety of both the plant, and the vessel and its crew, certain parts of the processing plant should be kept separate from other less dangerous process units. For simple material handling, the units should be roughly in adjacent sequence as the process steps, and utilize gravity flow whenever practical. The subdivision of the vessels and multiple deck levels permit both process unit separation and locations in appropriate proximity. 10.5.1.2 Nodule Storage The process plants would need a buffer stack of nodules on-board to assure continuity of production if the ship should become separated from the stowage or mining ship. The size of this nodule stowage hold would depend upon the mining and processing rates and the estimates of time out of connection. The postulated mining .rates are 3,750 and 12,500 tons per day, and two days separation may be expected if the mining ship loses its pipeline or because of storms. If nodule stowage is on a separate vessel from the process plant, the separa- tion problem because of weather becomes most significant, and a train of mining ship, nodule barge, and process ships is even more difficult to control. 10.5.1.3 Reagents and Supplies Storage Liquids and solids needed in processing include gaseous fuels in pressure vessels, acids, flocculants, chlorine and ammonia, and large volumes of raffinate from the shore plant for the pregnant liquor production cases. 10-20 These materials are used in large volumes, and a few are very difficult to handle and stow. Therefore, special facilities are required. 10.5.1.4 Steam All the at-sea processing plants require substantial flows of saturated steam which are not condensed but are consumed in the process. The basic shipboard steam plant includes feed water treatment, fuel oil treat- ment, water-wall boilers with pressures and temperatures in the 600 to 950 psi and 850 to 1050°F range, all superheated for efficiency. Modern steam turbines are in the 200 to 500 rpm region, with double reduction gears for the high, intermediate, and low-pressure stages. Reheat turbine plants are being adopted to improve efficiency of steam plants to be more competitive with large, slow-speed diesels for propulsion. Exhaust steam is condensed by sea water cooling and recycled through the system with relatively low losses. The process system requirements for saturated steam and electrical power would be met by either installing desuperheaters in marine boilers, or by using the superheated steam to generate electricity in the high-pressure turbine generator stages only. 10.5.1.5 Electrical Power Electrical power for the process plants must be generated on-board the pro- cessing vessel, as cable transfer at sea of the magnitude required would be too hazardous. Auxiliary electrical power is normally generated on ships by two or three steam turbine generators with their own condensers, and sometimes an in-port (small) boiler. This power may be needed for cargo handling equipment such as cargo oil pumps and heating, and electrical winches and cranes, as well as hotel services for the crew and power for propulsion system circulatory, treatment and cooling pumps, and fans. At most, auxiliary power generated is less than half of the propulsion plant powers for bulk ships and tankers, as listed in the transportation section. The electrical power requirements of the processes are somewhat in excess of normal maximum boiler output expected in conventional bulk ships. In almost all cases, however, sufficient power for process use can be obtained by letting high pressure superheated steam through backpressure turbines with the exhaust steam used for process heating. 10.5.1.6 Water Requirements A major addition for a process plant at sea would be installation of a large, sea-water evaporator or distiller to produce both the process and boiler make-up water. Normally, ocean-going ships with normal size crews do not distill water but carry fresh water from shore in tanks because of the relatively high cost of producing fresh water at sea and the low tonnage of water required. However, the volume of water required in all cases is far too large to transport and must be produced at sea. The sizes of the distiller plants are large enough to have efficiencies comparable to shore plants. 10-21 10.5.1.7 Hotel Facilities The hotel installation on board ship includes hot and cold fresh water, sewage, air conditioning, heating and ventilation, and sea water wash services- Standard galley, mess, cleaning, workshop, recreation, and other facilities are ordinarily provided. All of these services would be extended to the operators of the at-sea processing plant. Barges have none of these accommodations, and all would be required on a processing vessel. 10.5.1.8 Fuel Residual fuel oil could be burned in the boilers as fuel for power, steam, and fresh water. Large quantities are required, and could be carried either by the transportation ship on each load from shore, or separate tankers from Indonesia or Hawaii. Generally the separate tanker load would be in larger quantities, 35,000 to 100,000 ton loads. Again, large volumes of fuel bunker space and fast fuel oil transfer systems on the processing vessel would be required. Obviously, the plant aboard ship requires a careful re-analysis to reflect differing conditions from those normally assumed ashore for process plants. The supply of fresh water is quite limited and expensive at sea because of the fuel requirements. The need for saturated steam can be met by running all process superheated steam first through power turbines and condensing little recycled steam. Distiller water could be superheated directly so that wasting heat in condensers is limited. On the other hand, adequate cooling water is available in the ocean at little expense compared to land cooling methods. 10.5.1.9 Transfer to Other Vessels To handle the fuels, and raffinate and reagents from shore, and the pregnant liquor to shore, large pumping systems will be needed for the liquids, including transf er-at-sea hoses. For the solids which move in the lower volumes, container or pallet movement systems would be necessary, but these are cumbersome and dangerous. Many of the solids could be carried by the processing vessel in large enough quantities to last a year and eliminate transfer of solids except for foods. Transfer of the nodules from mining ship to process vessel is also required. 10.5.1.10 Feasibility of Processing Plants on Barges If processing plants were to be installed on a barge and then towed by tug- boat or the mining ship, a propulsion system would not be needed on the pro- cessing barge. However, boilers, distillers, generators, condensers, and all the other typical shipboard equipment would still be required on the process barge, as well as hotel fittings needed for the process plant crew. The barge would be manned at sea and have the same requirements for safety as a ship. Therefore, little gain would be possible by not installing propulsion and navi- gation equipment. In fact, because of the plant crew onboard, a standby tugboat may well be required at all times for emergency rescue, as is now common at the North Sea oil platforms. A barge processing plant could be moored at or near the mine site. In 15,000 feet of water, this is feasible for small units but is difficult and expensive. For multiple small barges or one large processing and stowage barge, 10-22 the deep-water mooring would be quite expensive and insecure in bad weather. If mining proceeds at 2 knots, a path 48 nautical miles long will be swept in one day. Therefore, a mining ship feeding a processing barge would need a long connecting floating pipeline. No doubt rather frequent repositioning would be required. For all of these reasons, a self-propelled, single ship processing plant appears to be a much more logical choice than a barge plant or several barges with plant and stowage spread between them. 10.5.2 Shipboard Process Plants This section reports on each of the five at-sea partial process plants on the vessel. 10.5.2.1 Reduction-Ammonia Leach The shipboard at-sea partial process plant equipment for the reduction/ ammonia leach process was listed in Table 10.1, and the services profile on Tables 10.2 and 10.3. Even when allowance is made for access space, installation room, separation of process segments, and ship subdivision with bulkheads, modest size vessels are adequate for these plants. A ship with 90-foot beam and less than 560 feet long, with a hull perhaps 50 feet deep, would have adequate space for all the process equipment, ship propulsion and hotel systems, distillers and generators, accommodations for a very large crew, lifesaving equipment, boats, and a heli- copter deck. Almost 12,000 tons of process equipment could be installed on a ship of the size indicated above. The vessel process pljnt inputs and outputs are tabulated in Table 10.12 for the pregnant liquor production case and in Table 10.13 for BMC productions. The inputs to the at-sea plant, based on 12,500 tons of nodules per day, total 16,400 tons of supplies, of which 13% is fuel, and 86% is raffinate returned from the land plant. The amount of nodule stowage, pregnant liquor, raffinate and fuel tankage, is a function of the frequency and schedule dependability of transporting ships to shore. If four Panamax ships of 55,000 DWT are used, a vessel would arrive every 2.7 days on the average. With a 24 hour leeway in arrival, a 63,000 ton storage capacity would be required for supplies. The pregnant liquor product to be shipped ashore would have separate tank capacity for 35,000 tons of liquid. Therefore the total weight of fuel and liquids for 3.7 days, and nodules for one day, is approximately 108,000 tons. The ship needed for both process plant and stowage would be about the same size as a 130,000 DWT tanker, that is 900'length, 138' beam and 53' draft when full laden at sea. This processing and stowage vessel would resemble a large tanker with some machinery near the center of the ship, on deck and in the holds, and would be a very stable platform. If larger size transport ships are utilized to call less frequently, even more stowage capacity would be needed. If three transport ships of about 66,000 DWT were used, about 15,000 more tons stowage capacity would be needed. 10-23 Table 10.12 Process Plant Inputs and Outputs At-Sea Partial Processing for Production of Pregnant Liquor Process Inputs Manganese Nodules (wet tons per day) Electrical Power (kilowatts) Saturated Steam (103 lbs per hour) Cooling Water (103 gals per min) Process Fresh Water (gal per min) Fuel Oil - Process Fuel (barrels Ships Power per day) Distiller Total Reduction Reduction Hi-Temperature Cuprion NH3 HCl H2SO4 NH3 Leach Leach Leach Leach 12,500 3,750 12,500 12,500 3-metal 4-metal 3-metal 3-metal 18,400 11,600 10,400 29,000 401 840 383 450 27 96 30 78 1,300 980 800 10,600 7,200 2,300 750 470 415 3,250 2,450 2,000 14,600 10,120 4,715 910 5,600 1,190 2,280 9,070 Propane or Butane (barrels per day) 1,400 Liquid Chlorine — 1,250 — (tons per day) Ammonia 12.1 — — (tons per day) Sulfuric Acid — — 255 (gals per min) Returned Tails Wash Liquor — — — (gals per min) Returned Reduction Liquor — — — (gals per min) Raffinate from Shore 2,360 — 3,550 (gals per min) Flocculants 3. 0.2 3 (tons per day) Total Inputs: Fuels (tons per day) 2,180 1,510 705 Supplies (tons per day) 15 1,440 Liquids (10^ gal per day) 3.40 — 5.11 Liquids (tons per day) 14,170 — 23,450 Total Input (tons/day) 16,370 2,950 26,950 Input Weight 1.31 0.79 2.16 13.3 3,500 1,620 1,350 16 7.37 30,730 32,100 2.57 10-24 Table 10.12 (continued) Process Inputs Output to Shore Pregnant Liquor (gals per min) (tons per day) Ratio: Output Weight Nodule Weight Outputs Overboard Tailings (tons per day) Reduction Reduction Hi-Temperature Cuprion NH3 HCl H2SO4 NH3 Leach Leach Leach Leach 1,350 8,100 0.65 8,180 557 4,500 1.20 805 3,550 23,400 1.87 10,400 3,830 23,000 1.84 10,500 Liquor (gals per min) 1,690 400 1,730 1,910 10-25 Table 10.13 Process Plant Inputs and Outputs At-Sea Partial Processing with Reduction of Metals from Pregnant Liquors Process Inputs Manganese Nodules (wet tons per day) Electrical Power (kilowatts) Saturated Steam (103 lbs per hour) Cooling Water (103 gals per min) Process Fresh Water (gal per min) Fuel Oil - Heat (barrels Power per day) Distiller Total Propane or Butane (barrels per day) Ammonia (tons per day) Sulfuric Acid (gals per min) Sulfur (tons per day) Lime (tons per day) Limestone (tons per day) Flocculants (tons per day) Total Inputs : Fuels (tons per day) Supplies (tons per day) Total Input (tons/day) Input Weight Ratio: ' ^, J 1 r, ° 1 Nodule Weight Output to Shore Precipitated Metals (tons per day) Output Weight Nodule Weight Ratio: Outputs Overboard Tailings (tons per day) Liquor (gals per min) Reduction Hi-Temperature Cuprion NH3 H2SO4 NH3 Leach Leach 12,500 Leach 12,500 12,500 3-metal 3-metal 3-metal 19,800 14,000 36,700 671 475 1,200 54 36 146 2,400 1,200 2,300 11,800 3,260 9,600 6,000 3,000 5,750 18,550 12.1 2,770 15 2,785 0.22 950 0.08 8,180 1,690 6,675 185 255 126 137 306 3 1,000 3,430 4,430 0.35 850 0.07 11,260 1,870 16,540 13.3 2,464 16 2,480 0.20 950 0.08 10,500 1,910 10-26 The hull cost of the processing plant-stowage vessel would be at least $50 million, and all of the processing plant and accommodations would be extra, perhaps running to as much as $200 million total for the large capacity ammonia- process vessel. Insurance would be difficult to get in present commercial markets; and only if a special effort were undertaken would such a vessel be fully insured in the future. The risks of explosion of the vessel and its large size would result in very high premiums, at least initially. If the concept of producing pregnant liquor at sea is used, then four transport tankers of 55,000 DWT would be needed. This compares with a need for only two ships of 63,000 DWT for slurry transport of raw nodules only. If the pregnant liquor is stripped at sea to produce a basic metal car- bonate cake, the output to the shore plant is reduced substantially, to about 8% of the weight of raw nodules, so that significant transportation savings will be possible. In addition, since no raffinate is returned, resupply tonnages decrease to less than 3,000 tons/day and consist mostly of fuel. Thus, a tanker could refuel the plant every two weeks, or even once a month, depending on pro- cess plant ship stowage capacity. A specially equipped general cargo ship could remove the basic metal carbonate cake at about the same interval. The vessel component dedicated to processing would have to increase to about 20,000 DWT equivalent to hold the additional equipment, but the overall size would decrease since less than 50,000 tons of fuel, 15,000 tons of cake and a days supply of nodules would be on board at any time. Therefore, the process vessel would have a size equivalent to a 100,000 DWT tanker. The transport ships would consist of one 40,000 DWT tanker and one 15,000 DWT cargo ship cost- ing about $30 million each. 10.5.2.2 High Temperature Sulfuric Acid Process The at-sea process plant units for the H2S0^ process are listed in Table 10.4. This plant, for pregnant liquor production, is about two-thirds the size of the ammonia plant of the same throughput capacity and smaller than the HCl . four-metal plant at one-third the capacity. A ship of about 15,000 DWT capacity could carry the plant, but not stowage of associated liquids and solids. Tables 10.12 and 10.13 summarize the at-sea plant inputs and outputs. The total weight of supplies is about 2.2 and the weight of pregnant liquors is 1.9 times the nodule weight. Of the order of four 70,000 DWT tanker ships, costing nearly $280 millions, calling about every 2.5 days, would be needed to handle the supplies and fuel input and raffinate out, which are about equal in tonnage. With a 24-hour cushion, and one days supply of nodules, about 190,000 tons of stowage capacity would be needed, and the complete ship would be over 200,000 DWT size. The hull and ship machinery would cost about $65 million, and the at-sea plant well over $100 million more. The ship would be over 1,000' long, 156' beam, 85' hull depth and draw 65'. If the pregnant liquor is treated at sea to produce metal sulfide preci- pitates, the plant size would be significantly increased and would approach the size required for the ammonia process, approximately 23,000 DWT. The product tonnage decreases to about 7% of the raw nodules weight, but a resupply of about 4500 tons/day is significant. The major input is s.ulibvic i^cid, which 10-27 under IMCO classification requires a moderate degree of containment, and at a rate high enough to require continuous resupply. Other supplies could be brought out monthly in a 15,000 DWT ship which would also return the sulfides to the shore plant. A 35,000 DWT tanker calling once a month could carry the required fuel and might also carry the propane/butane required for H2S production. However, capacity for 750 tons of LPC7 is an unusual requirement and pressurized tanks and gas recompression/refrigeration would be needed and the fuel oil/LPG carrier might cost about $30 million. The sulfuric acid would be transported in a chemical carrier, with special tank coatings and handling methods, along with the sulfur. A West coast sulfuric acid supply point would require a 40,000 DWT ship while a Gulf coast supply point would require at least one Panamax ship at a cost of $30-35 million. A single 15,000 DWT ship calling every two weeks would be adequate for the balance of supply /product transportation requirements and would cost about $30 million. The processing vessel itself would require stowage capacity for about 105,000 tons of supplies which, together with the process plant space, dictates a ship of about 130,000 DWT capacity. It would be about 850' long, 138' beam, and be 72' deep. 10.5.2.3 Reduction Hydrochloric Acid Leach The shipboard processing plant equipment for the HCl partial process is listed in Table 10.7. Although only 3,750 tons/day of nodules are processed in this four-metal plant, the plant equipment requires about 3/4 the space of the reduction/ammonia process. This plant should fit in a ship space two decks high and 400' x 80'. In addition, boiler, distiller, propulsion and generating plant space would be needed. Therefore, a ship of about 510' x 70' beam, or 17,000 DWT, could accommodate the plant, but have no storage space. Table 10.12 summarizes the at-sea plant inputs and outputs. This process does not require recycling of raffinate or liquor from shore, but does consume relatively more supplies at sea. This includes transportation and handling of large amounts of liquid chlorine. The output to shore in pregnant liquor is about 20% higher than the weight of nodules mined. Because of the lower tonnage involved, one tanker of about 70,000 DWT would be adequate for chlorine, fuel oil, and pregnant liquor movement to and from Southerm California ports. However, almost 13 days between ship calls would require about two weeks stowage capacity, or 42,000 tons of supplies and 63,000 tons of product, plus one day of nodules, 3 ,750 tons. Therefore, the stowage capacity required, approximately 96,000 tons, is the same range as for the ammonia process ship. Thus, the smaller, four -metal process plant makes little difference in the overall stowage and process plant vessel size, unless a larger fleet of smaller transport ships were used. The carriage of chlorine at sea is most hazardous, and requires a Class I type ship with separate cargo tanks, not only for the chlorine but all other chemicals onboard. At present the United States Coast Guard will not certify a manned vessel to carry chlorine in bulk quantities. However, about 200 tons per day are required by the HCl process plant, too much for barrel handling, and too little for efficient ship loads. If the chlorine cannot be carried in inde- pendent deck tanks on the tanker, then a barge load could be pulled. Presumably the process ship would be allowed to pump chlorine into holding tanks. 10-28 Carriage of both LPG and chlorine at the same time is not likely, although the same cryogenic cargo tanks and pumping system could be utilized. Therefore, the use of two transport tank ships costing about $33 million each, is antici- pated, one for LPG and the other for chlorine, and both carry fuel and floccu- lants. If the smaller four-metal HCl plant is installed at sea, the 90,000 DWT ship hull may cost approximately $30 million and the plant another $100 million installed, or about half of the three-metal, higher capacity plants. Even this, however, would exceed the cost of the most expensive LNG and tanker ships being built today in the United States, which are of comparable size. 10.5.2.4 Cuprion Ammonia Leach Process The Cuprion process at-sea plant equipment requirements are shown in Table 10.9. This plant is about as large as the ammonia reduction plant of the same throughput, and has many of the same components and processes. However, the process inputs and outputs are much larger for the pregnant liquor production case, as shown in Table 10.12. Inputs and outputs for BMC production are summarized in Table 10.13. Although the fuel oil required is less, the addition of returned raffinate and reduction liquors increases the daily outbound load from shore to process ship. Even the inbound load to shore is almost twice the mineral weight of manganese nodules. One large transport ship would call every other day, and six ships would be needed to transport the pregnant liquor ashore. Almost 10 million tons would be moved outbound annually. Again, the stowage capacity needed between ship arrivals, and the process plant together would require a large ship well over 215,000 DWT, which is comparable to the H2S0^ process ship in size. The dimensions of the hull would be about 1050' long, 160' beam, 88' depth, and 67' draft and the cost would exceed $70 million. As is the case with the reduction/ammoniacal leach process, pregnant liquor stripping for BMC production greatly decreases transportation requirements. With about 2500 tons/day resupply required, a small tanker calling once or twice a month would be adequate, depending on process ship stowage capacity. The same rate of call would be required for removal of BMC cake, but a separate ship would be needed and each would cost about $30 million. The process space required for BMC production would be less than that provided by a 20,000 DWT ship. Provision for stowage of fuel, supplies, and nodules, however, would result in a ship with a capacity of about 100,000 DOT being used. 10.5.3 Summary and Possible Alternatives to Selected Process Vessels 10.5.3.1 Summary In all four processes for the production of pregnant liquor, a large volume of liquid would be moved inbound to shore, in tanks. The largest item moved out to sea is raffinate (except for the HCl process) . In addition, the fuel oil transport tonnage alone would be substantial. However, this could be performed separately by tankers from Indonesia or Hawaii, rather than loaded out from the United States. Finally, supplies, in some cases hazardous liquids, and a small 10-29 cargo of dry chemicals would have to be transported. Therefore, tankers would be the most appropriate type of transport ships. In aggregate, the processes need enormous throughputs, summarized in Table 10.14, on a weekly basis to maintain a 12,500 ton per day nodule mining rate. Table 10.14 Weekly Tonnage for At-Sea Processing (short tons) Pregnant Liquor Production NH3 H2SO4 HC1>^ Cuprion Outbound Inbound 114,600 56,700 188,700 164,00 20,700 31,500 224,700 161,000 Total 171,300 352,700 52,200 385,700 Four-metal plant; throughput 30% of other capacities. The number of tanker ships to transport reagents and supplies ranges from one to six ships of very large size for the shortest haul. For longer hauls, a 50% to 200% increase in ships would be needed. Increases in costs would be of comparable magnitude. This large number of tankers would, of course, call frequently at the processing ship, with about every other day arrivals or every third day under ideal circumstances for the largest ship. The mining ship would probably begin receiving fuel and liquid supplies and raffinate or liquor for many hours of pumping operation before beginning to pump pregnant liquor to the transport tank ship. To allow for schedule uncertainties and maintain processing of nodules, a generous tankage of at least one-half week's operation would be desired at the ocean processing site. Therefore, the total tonnage listed on Table 10.14 would be double the necessary capacity of the processing ship. This is cargo and is in addition to in-process fluids and the process equipment weight. Therefore, the ship size needed for the process plant is moderate, as shown by the ammonia plant example. However, the ship size required for liquids stowage when producing pregnant liquor is quite large, between 90,000 and 215,000 DWT. Consequently, an integrated process plant and liquid stowage vessel will be enormous, about 1,000 feet in length, 150-foot beam, and 50-foot draft when normally loaded at sea. The production of precipitates holding the metal values eliminates the movement of pregnant liquor and the raffinate return which generated most of the transportation requirement and the need for large stowage capacity to discharge and load the transport ships quickly. (This method of precipitation does not apply to the four-metal HCl process at sea.) The process vessel stowage ranges 10-30 from 35% to 100% less than with the liquids for transport. Also, the transport vessel requirement is reduced for the three suitable processes — ammonia, sul- furic acid and cuprion. Table 10.15 compares the vessel requirements for all alternatives of partial at-sea processing and transportation. For the pregnant liquor production, the HCl process appears to ha'^e vessel advantages. For metal cake production, the reduction/ammonia and cuprion process are almost equal in vessel require- ments, and even comparable to the HCl (four -metal) at-sea plant. All three processes utilize a general cargo ship to move the metal cake to shore, less than 1,000 tons per day. All three processes need fuel carried by a tanker frequently to reduce oil bunker space requirements. In addition, the H2SO4 process needs a chemical tanker to carry the sulfuric acid volume needed. The worst hazards in sea transportation are generated by the large volumes of chlorine, ammonia, propane and butane (LPG) and sulfuric acid, in that order. The HCl processes needs large quantities of LPG and chlorine; and the high tem- perature sulfuric acid process also uses sulfur, LPG in small amounts, lime and limestone. On balance, an at-sea process based on ammonia leaching appears to present the lowest transportation safety problem. Table 10.15 Summary of At-Sea Process Vessel and Transportation Ships Pregnant Liquor Production Vessel Descriptions Reduction/ Ammoniacal Leach Reduction/ Hydrochloric Acid Leach High Tem- perature Cuprion/ Ammoniacal H-;)S04 Leach Leach Process Plant, (DWT) 130,000 (1) No. of Transport Ships 4 Size, (DOT) 55,000 90,000 43,000 205,000 70,000 215,000 70,000 Production of Metal Precipitates Process Plant, (DWT) 100,000 (2) No. of Transport Ships 2 Size, (DOT) 40,000 15,000 N/A N/A N/A 130,000 .(3) 100,000 o(2) 40,000 35,000 35,000 + LPG 15,000 15,00 (1) (2) (3) Tankers. Tanker plus general cargo ship. Two tankers plus general cargo ship, 10-31 Partial at-sea processing of manganese nodules is technically possible, but may become feasible only if land processing plants face formidable obstacles for operation and disposal of wastes. This analysis presupposes that wastes could be discharged at sea; if this is not the case, there would be no advantage to partial at-sea processing. Full processing at sea of manganese nodules to metals would doubtless have severe technical difficulties. The desire to reduce the volume of material transported to shore is difficult to achieve with partial processing. Substantial increases in transport weight of semi-processed nodules and process supplies may about double the transport ship investment and operating cost. However, the importance of these costs must be measured relative to the cost of the shore-based operation, which are not yet defined. 10.5.3.2 Process Alternatives The at-sea processing options for the production of pregnant liquor are dominated by the problems in handling pregnant liquor and raffinate, and re- duction in these volumes by a factor of two would have enormously favorable consequences. For this to occur, however, the centrifuges which were substituted for thickeners would have to function more efficiently than has been assumed. If higher slurry densities were attained, less wash water would need to be intro- duced into the leaching circuit and more concentrated pregnant liquors might be obtained. For this study the pulp densities produced in the centrifuges and the centrifuge wash efficiencies were conservatively assumed to be the same as those assumed for the thickeners specified in the land plants. While one might reasonably assume that, in fact, centrifuge performance would be better than thickener performance, no experimental data are available to support either the magnitude of the relative improvement or even the absolute value appropriate for thickeners on nodule leach residues. Thus, demonstrations of acceptable centri- fuge performance would likely be a precondition to development of at-sea process- ing technology. In addition, an increase in pulp density would decrease water losses which occur when washed tailings are dumped over the side. Since the fuel required for distilling fresh water amounts to 40% of the total required in some cases, the savings could be significant. Some desalinization will always be required for boiler water make-up and other uses, but if sea water could be used directly without excessive corrosion problems being introduced, there would be significant incentive to do so. Another alternative addressed was the division of the benef iciation and several processing steps into discrete sections to be accomplished on several vessels at sea. The partial at-sea processing plants were derived from examin- ation of the integrated plant designed for land operation. Proper plant design calls for use of heat from exothermic reactions and waste heat from wherever available in processes to be used where needed. When plant splitting is begun, these energy-saving possibilities are sometimes eliminated. Also, the need for certain reclaiming or cleaning operations at many places in the process, such as ammonia recovery, requires two such plants when the integrated process is split. The at-sea process plants could be split more, and installed on more than one vessel, with penalties as described above. But the principal problem of the amount of materials used at sea and transported would be aggravated rather than reduced. In case of delays of movement of liquids from one vessel plant to the next, buffer storage would be needed on the supplied ship to maintain processing, as well as dump storage on the supplying ship to maintain its 10-32 processing. Since this is already a major difficulty, worsened conditions are undesirable; and therefore, further split of the plant is not likely. One vessel stowage alternative is to utilize a fleet of barges alongside the processing ship, each one with a different commodity — fuel, acid, ammonia, fresh water, raffinate, or pregnant liquor. If six barges were used for a large plant, minimum sizes of 15,000 DWT to 75,000 DWT may be appropriate for different liquids. However, efficient transfer of these materials at sea is a problem. Similarly, mooring and connection problems are difficult, especially when all must proceed at slow speed with the mining ship. The next stowage alternative is to utilize the transport ship or barges as the at-sea stowage tanks. Twice to three times per week the transport ship may be replaced from its alongside position and a new ship moved into transfer location until its cargo is consumed. These transport ships would then be closely connected to the processing ship. Relatively little stowage would be needed on the processing ship, perhaps only a half-days operation for fast-moving items. In this concept, the vessel compartments would switch slowly from the outbound cargo contents to the lesser inbound load. Because of the extended time the transport ship or barge is alongside the processing ship, the number of vessels would increase, if size cannot. 10.5.3.3 Vessels Alternatives The offshore processing plant may be constructed on new hulls, or instal- led on an existing vessel. A substantial number of large barges are in service under American flag. Although a detailed inventory, such as available from the Coast Guard or Customs, has not been analyzed, many large barges adequate in size for installation of a processing plant may be available. However, as noted above, the use of barges is considered unlikely for at-sea processing. Table 10.16 lists the ocean-going U.S. flat dry bulk ships in service as of January 1, 1977. Of the 19 ships, three have been built since World War II and may be useable. However, the two OBOs of Aeries Marine Shipping are the only large size ships, and they would be better as transport ships since they are not large enough to both process and stow the volumes of liquids and nodules. Table 10.17 lists the large U.S. flag tankers over 51,000 DWT as of Janua- ry 1, 1977. These vessels are large enough to convert to processing ship service, if desired. However, the propulsion plants are of marginal power and the accommodations inadequate. A new hotel, probably forward with navigating spaces, would be required in addition to the installation of processing plant, subdivision of cargo tanks, and related work. Therefore, a vessel processing plant would probably be built new. 10.6 CHARACTERIZATION OF LAND-BASED PROCESSING PLANTS FOR AT-SEA PREPROCESSING 10.6.1 Reduction/Ammoniacal Leach Process The separation between at-sea and land-based processing for the Reduction/ Ammoniacal Leach Process is shown in Figure 10.2. The configuration of the land- based plant for pregnant liquor processing is essentially identical to that described in the base case (full processing on land), except that the front-end, 10-33 Table 10.16 American Flag Ships Available * Name Ultramar, Ultrasea Bethflor, Bethtex Rice Queen Marine Electric Overseas Traveler Kopaa Albany, James, Mohawk, Yellowstone Columbia, Plattes, Potomac, Merimac Sugar Islander Inges, Walter Rice Type DWT Horsepower Built OBO 82,100 T-24,000 1975-76 Bulk 23,500 T-13,000 1948 Bulk 14,400 T-7240 1944 Bulk 25,600 T-7240 1944 Bulk 24,100 T-7240 1945 Bulk 24,200 T-10,000 1944 Bulk 15,500 T-9900 1945 Bulk 26,200 T-7240 1945 Bulk 29,600 D-12,000 1973 Bulk 23,500 T-7240 1945 Bulk ships suitable for ocean ore carriage with modifications Source: Marine Engineering/Log Magazines 10-34 Table 10.17 Large U.S. Flag Tankers (Over 51,000 DWT) Name Massachusetts, New York Mobil Arctic Prince William Sound Arco Anchorage, Juneau, Fairbanks Overseas Juneau Manhattan Golden Dolphin, Endeavor, Monarch, Worth, Beaver State, Rose City Joseph D. Potts, Sohio Intrepid, Resolute, American Sun Exxon San Francisco, Philadelphia, Baton Rouge Vantage Defender Sansienena II Arco Prudhoe Bay, Sag River Chevron California, Mississippi, Hawaii Exxon Houston, New Orleans Overseas Arctic, Alaska, Golden Gate Baltimore Trader Exxon-Boston DWT Horsepower 35,000 Built 265,000 1976 129,000 30,000 1972 124,000 30,000 1976 120,600 26,000 1973-74 120,500 26,000 1973 114,700 43,000 1962 91,800 24,500 1974 81,000 24,000 1969-71 75,600 19,000 1970 71,000 25,000 1976 70,500 20,000 1971 70,400 20,000 1971-72 70,200 20,000 1972-73 67,900 19,000 1964-65 62,000 1973 57,900 15,000 1955 52,000 19,000 1960 10-35 consisting of nodules receiving, ore preparation and drying, reduction, leaching, and washing, as well as producer gas generation, are not included. In addition, the tailings stripping operation is removed from the ammonia recovery section. The land-based plant then consists of the liquid ion exchange extraction and stripping sections, copper and nickel electrowinning, cobalt recovery, ammonia recovery, and materials handling and services sections. If a mixed basic metal carbonate (BMC) is produced at sea and processed in a land plant, additional processing steps are required for BMC dissolution and carbonate removal. The former is carried out by dissolving the wet BMC cake in recycled raffinate while the carbonate is removed by reactions with lime. The plant inputs and outputs are summarized in Tables 10.18 and 10.19. The requirement for raw materials, particularly coal, is greatly reduced because reduction and the tailing stripping occur at sea. Purchased power, however, is not greatly reduced because the electrowinning operations still occur on land. The raw nodules input is replaced either by a large flow of pregnant liquor (and an additional output, raffinate, is produced which must be returned to the at-sea pre-processing plant) or by a small flow of BMC (with nothing returned) to sea. The amount of tailings and wastes produced is greatly reduced, consisting in this case mainly of solids from the lime boil operations. The requirement for support services, particularly transportation, will be reduced in proportion to the input of production supplies. The elimination of the front-end operations will reduce the labor requirements somewhat. It is estimated that about 400-450 people will be employed in the land-plant. In addition, elimination of the leach and tailings wash thickeners and the require- ment for a large raw nodules storage area would decrease the plant area require- ment to approximately 150 acres. 10.6.2 High Temperature Sulfuric Acid Leach Process The separation of at-sea and land-base processing operations for the high temperature sulfuric acid leach process is shown in Figure 10.3, The configur- ation of the land-based plant for processing pregnant liquor is identical to that of the base case, with the exception that nodules receiving, ore prepara- tion, leaching, and nodules washing are carried out at sea. The land-based plant consists of pH adjustment and raffinate neutralization steps, copper and nickel liquid ion exchange extraction, stripping, and electrowinning steps, cobalt recovery, ammonia recovery, and materials handling and services operations. If a mixed metal sulfide precipitate is produced at sea for processing in a land-based plant, an additional step, sulfide dissolution, is required. This is accomplished by oxygen pressure leaching in a manner similar to that used in the smelting process. All other process steps are the same as for pregnant liquor processing, except that the materials requirement for the pH adjustment and neutralization steps are reduced because the leach solution is more concen- trated in metals and largely free of other impurities. 10-36 Table 10.18 Land Plant Inputs and Outputs Reduction/Ammoniacal Leach Process At-Sea Processing Pregnant Liquor Production Stream No . Composition Rate Products 93 112 143 144 138 Major Inputs 32 182 187 188 200 Cathode Copper 10^ tons/yr Cathode Nickel 10-^ tons/yr Nickel Powder 10-^ tons/yr Cobalt Powder 10^ tons/yr Zinc/Copper Sulfide 10^ tons/yr Pregnant Liquor 10" gal/yr Coal 103 tons/yr Limestone 10^ tons/yr Lime 10^ tons/yr Water 10^ gal/yr Purchased Power 10^ Kwh 30.4 35.6 0.7 4.3 2.4 642 53.5 1.6 9.3 713 183 Major Outputs 41 161 161 225 154, 186 97, 120 204 Return Raffinate Liquid 10 gal/yr Solid Wastes 10^ tons/yr Liquid Wastes 10-^ gal/yr Stacked Off-gases 10-^ std cu ft/min Low Level Emissions: Vents 10-^ std cu ft/min Electrowinning Vents 10-^ std cu ft/min Cooling Tower Evaporation 10^ std cu ft/min Production Materials and Supplies Gases 172, 191 60, 113, 134, 137 140 207 Liquids NH3 10^ tons/yr H2S 103 tons/yr H2 106 std cu ft/ yr CI2 10^ tons/yr 76, 96, 117, 142, 119 H2SO4 10^ gal/yr 145, 194 Fuels (POL) 10^ gal/yr 72 LIX Reagents 10^ gal/yr Solids 43 95 115 116 118 202 Flocculants 10^ tons/yr Electrowinning Additives tons/yr Na2S04 tons/yr H3BO3 tons/yr Carbon tons/yr Tower Water Treatment tons/yr 1121 28. 32. 67 50 9. 15 75 4.0 4,9 96 0.1 670 500 75 0. 16 1350 200 40 27 10-37 Table 10.19 Land Plant Inputs and Outputs Reduction/Amraoniacal Leach Process At-Sea Processing Basic Metal Carbonate Production Stream No , Composition Rate Products 93 112 143 144 138 Major Inputs 32 182 187 188 200 Cathode Copper 10^ tons/yr Cathode Nickel 10-^ tons/yr Nickel Powder 10-^ tons/yr Cobalt Powder 10^ tons/yr Zinc/Copper Sulfide 10-^ tons/hr BMC Cake 10 tons/yr Coal 10^ tons/yr Limestone 10^ tons/yr Lime lo3 tons/yr Water 10^ gal/yr Purchased Power 10^ Kwh 30.4 35.6 0.7 4.3 2.4 313 53.5 1.6 49.7 190 188.5 Major Outputs 161 Solid Wastes 10^ tons/yr 161 Liquid Wastes 10^ gal/yr 225 Stacked Off-gases 10^ std cu ft/min Low Level Emissions: 154, 186 Vents 10^ std cu ft/min 97, 120 Electrowinning Vents 10^ std cu ft/min 204 Cooling Tower Evaporation 10-^ std cu ft/min Production Materials and Supplies Gases 172, 191 NH3 10^ tons/yr 60, 113, 134, 137 H2S 103 tons/yr 140 H2 10^ std cu ft/ yr , 207 CI2 103 tons/yr Liquids 76, 96, 117, 142, 119 145, 194 72 H2SO4 10^ gal/yr Fuels (POL) 103 gal/yr LIX Reagents 103 gal/yr 100.8 49.6 67 50 9. 17 75 4.0 4.9 96 0.1 670 500 75 Solids 43 95 115 116 118 202 Flocculants 10^ tons/yr Electrowinning Additives tons/yr Na2S04 tons/yr H3BO3 tons/yr Carbon tons/yr Tower Water Treatment tons/yr 0, 16 1350 200 40 27 10-38 Plant inputs and outputs are summarized in Tables 10.20 and 10.21. While the steam requiremenLs for leaching are eliminated, those for ammonia recovery remain the same and a significant amount of coal is still required. The pur- chased power requirements remain high since the electrowinning power demands are only partially offset by internal power generation. The consumption of sulfuric acid is greatly reduced, since leaching takes place at sea, and raw nodules input is replaced by a large flow of pregnant liquor. While process solid wastes are decreased considerably since tailings are disposed of at sea, the lime-boil operation still produces a significant amount of gypsum which, in this case, constitutes approximately 90 to 95% of the total wastes discharged. The requirement for support services is also diminished because of the decreased flow of raw materials, and elimination of the front-end operations would reduce labor requirements to approximately 400'-450 people. Elimination of tailings washing thickeners and raw nodules storage areas would reduce plant land requirements to approximately 130 acres. 10.6.3 Reduction/Hydrochloric Acid Leach Process The separation between at-sea preprocessing and land-based operations for the reduction/hydrochloric acid leach process is shown in Figure 10.4. The process configuration is essentially the same as for the base case, except that nodules receiving, preparation, and drying, hydrochlorination and hydrolysis and leaching and washing are carried out at sea. Hydrogen chloride and chlorine recovery are carried out both at sea and on land, and a chlorine burning (hydrogen chloride production) step is added to the at-sea processing sequence. Copper and nickel ion exchange extraction and stripping and electrowinning are carried out in the land plant as are cobalt extraction and stripping and cobalt recovery, chloride salt drying and fused salt electrolysis for manganese recovery, and the provision of process materials handling and services for the land-based operations. The chlorine recovered in the land-based plant is com- pressed and liquified for return to the at-sea preprocessing plant. Elimination of the requirement for large volumes of hydrogen chloride for use in the land- based plant eliminates the need to tie this facility to an adjacent caustic- chlorine/chlorinated hydrocarbon production facility. However, the requirements of the at-sea plant are such that an additional 400 tons/day of chlorine would have to be purchased from external sources to supplement the chlorine recovered from the land-based operations. Plant inputs and outputs are summarized in Table 10.22. While the require- ment to trade chlorine and hydrogen chloride over-the-f ence is eliminated, purchased power needs remain high to support the electrolytic reduction of man- ganese. Coal consumption is reduced significantly since the large steam demand associated with hydrogen chloride recovery has been eliminated. While at-sea disposal of leach tailings reduces this component of solid waste significantly, the amount of fused salts to be disposed of remains the same since they are produced during the manganese reduction step. The requirement for plant support components would also decrease somewhat, with personnel levels dropping to 250 people because of the elimination of the front-end operations. Elimination of tailings washing and raw nodules storage areas would decrease total land requirements to approximately 75 acres, and the 10-39 Table 10.20 Land Plant Inputs and Outputs High Temperature Sulfuric Acid Leach Process At-Sea Processing Pregnant Liquor Production Composition Rat e 3 Cathode Copper 10 tons/yr 32 Cathode Nickel 103 tons/yr 38 Nickel Powder 10^ tons/yr 0.2 Cobalt Powder 103 tons/yr 4.1 Zinc/Copper Sulfide 103 tons/yr 5.2 Pregnant Liquor 10° gal/yr 1687 Coal 103 tons/yr 209 Limestone 10^ tons/yr 96 Lime 103 tons/yr 350 Water 106 gal/yr 782 Purchased Power 10^ Kwh 115. Sulfuric Acid 103 tons/yr 63 3 Return Raffinate Liquid 10 gal/yr 826 Solid Wastes 103 tons/yr 716 Liquid Wastes 10^ gal/yr 236 Stacked Off-gases 103 std cu ft/min 150 Low Level Emissions: Vents 103 std cu ft/min 55.5 Electrowinning Vents 103 std cu ft/min 11.63 243 Cooling Tower Evaporation 103 std cu ft/min 115 Production Materials and Supplies Gases 202 NH3 103 tons/yr 2.4 80, 174 H2S 103 tons/yr 7.3 180 H2 106 std cu ft/yr 89.3 252 Cl2 103 tons/yr 0.36 Liquids 92, 126, 144, 155 H2SO4 103 tons/yr 62 191, 177 183, 228 Fuels (POL) 103 gal/yr 552 LIX Reagents 103 gal/yr 247 Solids 38 Flocculants tons/yr 2.8 97 Electrowinning Additives tons/yr 16 153 Na2S04 tons/yr 610 154 H3BO3 tons/yr 120 152 Carbon tons/yr 40 Tower Water Treatment tons/yr 206 10-40 Stream No. Pro ducts 90 150 181 182 178 Ma-j 3T Inp uts 32 222 225 226 240 92, 126, 144, 155 177, 191 Ma,T or Out puts 31 207 207 248 193 , 210, 224 96 , 160 Table 10.21 Land Plant Inputs and Outputs High Temperature Sulfuric Acid Leach Process At-Sea Processing Mixed Metal Sulfide Production Stream No . Composition Rate Products 90 Cathode Copper 10 tons/yr 32 150 Cathode Nickel 10^ tons/yr 38 181 Nickel Powder 10^ tons/yr 0.2 182 Cobalt Powder 103 tons/yr 4.1 178 Zinc/Copper Sulfide 103 tons/yr 5.2 Major Inputs 32 Metal Sulfide Cake 10"^ tons/yr 274 222 Coal 103 tons/yr 60 225 Limestone 10^ tons/yr 60 226 Lime 103 tons/yr 96 240 Water 10^ gal/yr 499 Purchased Power 10^ Kwh 189 92, 126, 144, 155 Sulfuric Acid 400 177, 191 Major Outputs 207 Solid Wastes 10^ tons/yr 250 207 Liquid Wastes 10^ gal/yr 100 248 Stacked Off-gases 103 std cu ft/min 41 Low Level Emissions: 193, 210, 224 Vents 10^ std cu ft/min 55.5 96, 160 Electrowinning Vents 10^ std cu ft/min 11.63 243 Cooling Tower Evaporation 103 std cu ft/min 45 Production Materials and Supplies Gases 02 103 tons/yr 60 202 NH3 103 tons/yr 1.8 80, 174 H2S 103 tons/yr 4.0 180 H2 106 std cu ft/yr 100 252 CI2 103 tons/yr 0.36 Liquids 92, 126, 144, 155 191, 177 H2S0^ 103 tons/yr 400 183, 228 Fuels (POL) 103 gal/yr 552 LIX Reagents 103 gal/yr 160 Solids 38 Flocculants tons/yr 2. 97 Electrowinning Additives tons/yr " 16 153 Na2S04 tons/yr 610 154 H3BO3 tons/yr 120 152 Carbon tons/yr 40 Tower Water Treatment tons/yr 206 10-41 Table 10.22 Land Plant Inputs and Outputs Reduction/Hydrochloric Acid Leach Process At-Sea Processing Pregnant Liquor Production Stream No . Composition Rate Products 73 Cathode Copper 10^ tons/yr 9.6 142 Cathode Nickel 103 tons/yr 12 195 Nickel Powder 103 tons/yr 0.2 196 Cobalt Powder 10"^ tons/yr 1.9 190 Zinc/Copper Sulfide 103 tons/yr 1.6 163 Manganese Metal 103 tons/yr 206.8 Major Inputs 32 Pregnant Liquor 10° gal/yr 256 262 Coal 103 tons/yr 65 266 Limestone 103 tons/yr 3 267 Lime 103 tons/yr 4.2 280 Water 10^ gal/yr 190 Purchased Power 10^ Kwh 864 52, 91, 112 50% Caustic Solution (as NaOH) 103 tons/yr 29.7 Major Outputs 173 Fused Salts 10^ tons/yr 233 222 Chlorine 103 tons/yr 413 243 Solid Wastes 103 tons/yr 17.5 243 Liquid Wastes 10^ gal/yr 16.6 300 Stacked Off-gases 103 std cu ft/min 80 Low Level Emissions: 205, 265 Vents 103 std cu ft/min 52.3 77, 150 Electrowinning Vents 103 std cu ft/min 3.46 284 Cooling Tower Evaporation 103 std cu ft/min 34 Production Materials and Supplies Gases ). 124. 143. H2S 103 tons/yr 1.8 NH3 10 tons/yr 50 N2 10 tons/yr 187 H2 106 std cu ft/yr " 41 CI2 103 tons/yr 0.1 H2SO4 10^ gal/yr 430 Fuel (POL) 103 eal/yr 500 LIX Reagents 103 gal/yr 57 176 Borax tons/yr 67 37 Flocculants tons/yr 4 75 Electrowinning Additives tons/yr 5 145 Na2S04 tons/yr 360 146 H3BO3 tons/yr 52 148 Carbon tons/yr 20 182 Electrodes tons/yr 400 10-42 Tower Water Treatment tons/yr 56 Boiler Water Treatment tons/yr 12 64, 99, 124, 143, 174, 206 245 175 192 287 Liquids 76, ,147, 194, 232 149 197, 270 53, 92, 113 Solids 282 289 transportation network to support plant activities would decrease in proportion to the requirements for raw materials. 10.6.4 Cuprion/Ammonia Leach Process The separation of at-sea and land-based processing operations for the Cuprion/Ammonia Leach Process is shown in Figure 10.5. The configuration of the land plant is the same as for base case conditions when pregnant liquor is processed except that nodules receiving, ore preparation, reduction, leaching and washing, producer gas generation and tailings stripping occur at sea. Liquid ion exchange extraction and stripping, copper and nickel electrowinning , cobalt recovery, and the raffinate stripping operations in ammonia recovery are retained in the land plant as are the materials handling and services sections. When processing a BMC produced at sea, however, the additional steps of cake dissolution and carbonate removal are required in the land plant. Also, since no liquors are returned to sea, the raffinate stripping operation is eliminated from the land plant as well, which greatly reduces steam generation and other service requirements. Plant inputs and outputs are summarized in Tables 10.23 and 10.24. The reduction in coal requirements is not as great as for the Reduction/Ammonia Leach Process when processing pregnant liquor, because the raffinate stripping step remains in the land-based plant in this configuration. The additional power generated from steam raised for this step gives a proportionately greater decrease in purchased power requirements for this case, however. As is the case for the Reduction/Ammonia Leach Process, the amount of process waste is significantly reduced since tailings are disposed of at sea, leaving mainly lime-boil solids and combustion ash to be disposed of on land. The raw nodules input is eliminated, being replaced by pregnant liquor, and a corresponding amount of raffinate is produced for return to the sea-based operation. Since raffinate stripping is performed at sea when processing BMC, the steam and coal requirements are reduced by amounts comparable to the Reduction/ Ammonia Leach case. And, of course, a large flow of pregnant liquor is replaced by a small amount of BMC cake and return of raffinate is eliminated. The requirements for support services are also reduced, as is the case for the other process options examined. Transportation requirements are decreased in proportion to the input of raw materials, manning levels decrease to approxi- mately 400-450 people because of the elimination of front-end processing requirements, and plant area requirements decrease to 150 acres because of the elimination of leaching and tailings washing thickeners, and raw nodules storage ponds. 10-43 Table 10.23 Stream No, Land Plant Inputs and Outputs Cuprion/Aimnoniacal Leach Process At-Sea Processing Pregnant Liquor Production Composition Rate Products 93 112 143 144 138 Cathode Copper 10^ tons/yr Cathode Nickel 10^ tons/yr Nickel Powder 10-^ tons/yr Cobalt Powder 10^ tons/yr Zinc/Copper Sulfide 10^ tons /yr 30.4 35.6 0.7 4.3 2.4 Major Inputs 32 182 186 187 200 Pregnant Liquor 10" gal/yr Coal 10^ tons/yr Limestone 10^ tons/yr Lime 10^ tons/yr Water 10^ gal/yr Purchased Power 10^ Kwh 1820 360 12. 9. 950 None Major Outputs 41 161 161 225 154, 185 97, 120 203 Return Raffinate & Reduction Liquor 10'^ gal/yr Solid Wastes 10^ tons/yr Liquid Wastes 10^ gal/yr Stacked Off-ga-3i'.s 10-^ std cu ft/min Low Level Emissions: Vents 10-^ std cu ft/min Electrowinning Vents 10^ std cu ft/min Cooling Tower Evaporation 10^ std cu ft/min 1958 125 125 250 55. 9. 150 3 75 Production Materials and Supplies Gases 12, 169 60, 113, 134, 137 191 Liquids 76, 96, 117, 119, 142 145, 194 72 Solids 43 95 115 116 118 193 19?. NH3 10^ tons/yr UyS 103 tons/yr CI2 103 tons/yr H2SO4 103 gal/yr Fuels (POL) 10^ gal/yr LIX Reagents 10^ gal/yr Flocculants 10^ tons/yr Electrowinning Additives tons/yr Na2S04 tons/yr H3BO3 tons/yr Carbon tons/yr NaCl tons/yr Tower Water Treatment 4.4 4.0 0.1 670 500 105 5 16 1350 200 40 260 270 10-44 Table 10.24 Land Plant Inputs and Outputs Cuprlon/Ammoniacal Leach Process At-Sea Processing Basic Metal Carbonate Production Stream No . Products 93 112 143 144 138 Composition Rate Cathode Copper 10^ tons/yr Cathode Nickel 10-* tons/yr Nickel Powder 10^ tons/yr Cobalt Powder 10^ tons/yr Zinc/Copper Sulfide 10^ tonsA^r- 30.4 35.6 0.7 4.3 2.4 Major Inputs 32 182 186 187 200 Major Outputs BMC Cake 10^ tons/yr Coal 10^ tons/yr Limestone 10^ tons/yr Lime 103 tons/yr Water 10^ gal/yr Purchased Power 106 Kwh 314 64 2.5 49.7 190 183 161 Solid Wastes 10^ tons/yr 153 161 Liquid Wastes 10^ gal/yr 138 225 Stacked Off-gases 10^ std cu ft/min 45 Low Level Emissions: 154, 185 Vents 10^ std cu ft/min 55. 97, 120 Electrowinning Vents 10^ std cu ft/rain 9. 203 Cooling Tower Evaporation 10^ std cu ft/min 19 Production Materials and Supplies Gases 12, 169 NH3 10^ tons/yr 60, 113, 134, 137 H2S 103 tons/yr 191 CI2 103 tons/yr 4.4 4.0 0.1 Liquids 76, 142 96, 117, 119 145, 72 Solids 43 95 115 116 118 193 192 194 H2SO4 103 gal/yr Fuels (POL) 10^ gal/yr LIX Reagents 10^ gal/yr Flocculants 10^ tons/yr Electrowinning Additives tons/yr Na2S04 tons/yr H3BO3 tons/yr Carbon tons/yr NaCl tons/yr Tower Water Treatment tons/yr 670 500 105 5 16 1350 200 40 260 270 10-45 11.0 WASTE TREATMENT 11.1 WASTE TREATMENT IN THE BASE CASE The base case process descriptions, including material and energy balances, were carried out on process configurations deemed most likely to be representative of those which would be employed in first generation nodule- processing plants. Required levels of emission controls were assumed; i.e., process vents must be contained and/or manifolded to scrubbing steps and process off-gases must be treated to ensure that emission control standards for particulates, gaseous sulfur emissions, and nitrogen oxide emissions are met. Likewise, adequate measures must be taken for process dust control, including the use of high-volume ventilation followed by scrubbing, the use of electrostatic precipitators, or bag houses as required, to return dusts to the process. The solid and liquid wastes produced during treatment of off- gases are combined with other process solid or liquid wastes and discharged. Prior to discharge, we assume the wastes would be treated only to obtain the desired pH and would be transported, in slurry form, to lined tailings ponds for ultimate containment. We have assumed that provisions would be made in the plant design to collect plant runoff and accidental spills, which would be mixed with other process solid and liquid wastes so that a single solid/ liquid effluent would be discharged from the plant. The process configurations were developed in response to perceived conventional technical and economical needs. For instance, it was not deemed necessary to evaporate cooling tower and boiler blowdown steams to achieve zero liquid discharge from the sources, since they are so small compared to the discharge of liquid wastes associated with tailings slurries, for instance, as to have negligible impact on the overall plant water balance. The use of techniques, such as centrifugation, instead of conventional thickening, to achieve liquid/solid separations in order to reduce the volume of waste, was not considered in view of the adverse economics which would likely result. In this section, alternative methods of waste handling will be investi- gated. The objective of this study is to describe and determine the impact that alternative methods of waste handling would have on the overall plant material and energy balances and on the characteristics of liquid and solid effluents. Since the original process descriptions assume that gaseous effluents are treated to the required degree, they will not be considered further here. 11.2 TYPES OF WASTES The solid and liquid effluents from the various nodules treating process- es are classified according to their chemical and physical characteristics in Tables 11.1 and 11.2. The streams within each process having the noted char- acteristic are defined. Question marks are used to denote cases where some uncertainty exists as to the appropriate classification of a particular stream. The rationale for the classification of a material as a toxic element has been covered previously. In this case, however, it is taken to include the heavy metals not so classified previously. Thus, tailings material which had passed through leaching and exhaustive washing steps could still be characterized 11-1 H XI: O ■H U Cf •H t-i 01 ■U CJ ta Vj crt x: u rH « o •H E CO ^ 0) o w o 01 u 4-1 pL, en (fl CO 3 a) .H U-l 3 o •n o c z o •H ■U cfl (J •H •H CO CO cfl ^ U _ c^- -i rH N C^. #. c^. M O r-\ I^J C^» C^ CM t-^ CM a. CO " CM -* * •* CM «VD m c^. >> CO rH ro •> O c- «. c^. Ofl o m C^' r-^ rH CN ^O c eg - vD CM CO CN -* •iH - iH CO - CO m CO 4-1 CSI C3N CO CO •* ». iH VO CM o- »* vO c^- c^- C^. o. 0) ^ - o C'- CN r- O rH ^ E - rH CM T3 hJ n #« CN ^ >. rH rH ^ r-{ pa 13 CO 00 O' 00 ^ -H CN CN CN (N C o •V *• in •» o < ^ c^- C^' Csl c^- •H -a- vD CN *• CN 4-1 O rH C3N in C^' in O -H •- CN CM ^— N O CN d u CN c^- •N C^. ffi .H '■ rH rH - in rH in d O u-1 -* O in ^ in en ro CN CN CO CM CN CM x: ^ ^ o CO •« CM CN ~^ CU n c^- •* ^"^ s— ^ C iJ t^ rH C^* C^* C^' o rH CN o in in •H .H - CN eg o o 4-1 CO O « CN CM (N CJ CJ vD c^' r^ " c^. ^ £>. 3 CO rH O C^' C^- rH rH rH •d -H - CN UO CO CM O CN 0) C . O -rl c O o o o c a. H xn pU M CO CO H CO M (U CU en o CU XI u > i-H .n o iH CM !J PM (U f^ iH ^ K ^ a) tfl w ^ H -i o > CNJ SC T) •• ^ -H CN C O U-1 r-t O < CN CO •H «i CM 4J O O »• O -H -a- ■- <>■ rH ^— ' 1 O o in O •s in vO vO rH CM >* o rH rH CN eg CJN 04 CO cu ,— s 60 CO 0) rH X) CO 0) en ^-- 0) 3 T3 x CO 01 0) 0) rH •rl CO -V o 4J s CO rH . >^ 0) •H o o C O H X CO > CJ> CO hH CO 0) 01 CO o •a cu cu (U 0) -g 3 c s cfl cu u CIH e B Fi 0) (fl cfl cn VJ (U 0) OJ Vj 1-1 u to 4-J 4-1 4-1 U CO CO CO CU XI x; X X H 4J 4-1 J-) -1 •H •H •rl z 3 3 3 11-3 as containing fixed toxic elements if, for instance, some of the minor toxic elements were not completely removed with the value metals during leaching. Not included are those materials, such as the precipitates obtained during water- softening operations which contain no toxic elements, or at least, for which these materials are present in negligible amounts. Coal combustion ashes and slags and stack-gas scrubber sludges might be classified in this category, depending upon the composition of the coal and the efficiency of particulate removal processes prior to scrubbing. Soluble solids are those materials which, though in solid form, will dissolve if exposed to the action of, say, rainwater over long periods of time. Most solutions in contact with process solid wastes in slurry form will contain at least some toxic elements. Solutions containing only innocuous compounds would be the liquid in contact with water-softening slurries and, depending upon the treatment regime used, cooling tower and boiler blow-down streams and boiler-feed water deionization purges. It is worth noting again, that unless a toxic element is incorporated into a product for sale, it must exit the process in a waste stream if it has been introduced in either the raw nodules or one of the process supplies. The specific treatment options considered, then, will either transform the toxic element once more to isolate it from accompanying wastes, isolate and change its chemical or physical form, or precipitate it in situ for disposal with other wastes in a less noxious form. Similar considerations apply to those innocuous constituents which are solubilized during the processing or introduced (such as sea salt) with raw nodules, and exit in mixed process waste streams. If soluble, innocuous compounds are not removed, they may, over extended periods, leach into ground water supplies when otherwise inert solid wastes are contacted with rainwater in disposal sites not properly lined. The classification of wastes by their physical characteristics, Table 11.2, is somewhat subjective in that the behavior of slurries, in particular, is difficult to define in the absence of operating data. The difference between hydrous slimes and settleable sludges is one of both water content and the ultimate stability of the material. Hydrous slimes, such as those formed during phosphate rock processing, for instance, do not increase greatly in slurry density upon standing for long periods of time, and are not easily stabilized by conventional techniques. Settleable sludges, on the other hand, would be amenable to stabilization by conventional techniques after de-watering and would probably support subsequent activity in the disposal area. Vitreous solids, such as smelting slags and slags from coal gasification, tend to be of relatively large particle size, de-water easily, are relatively inert chemi- cally, and can be disposed of relatively easily or even used for road-bed materials, as is the case with non-ferrous smelting slags. Combustion ash, on the other hand, is finer, must be wetted to prevent dusting, and is not as inert chemically. Ash is leachable on exposure to rain under certain circum- stances. The inert solids are characterized by a relatively large particle size and by the fact that they are obtained from the process essentially dry or with only surface moisture. They represent small waste streams which can be easily handled, except that they contain toxic elements in varying amounts. The wastes discharged in clear solutions may, or may not, be toxic, and could be treated by conventional means if not admixed with much larger volumes of other process solid wastes. 11-4 11.3 WASTE TREATMENT ALTERNATIVES The following waste treatment alternatives have been reviewed for their impact on the overall process material and energy balances: • Separate the various wastes in process and treat each one according to the requirements imposed by the nature of the particular waste and the requirements for the final physical form desired; • Modify the basic process configuration to either eliminate the occurrence of a particular toxic constituent or change the physical form of the waste material to conform with the desired final physical state, e.g., dried solids, slurry, etc. ; • Treat process wastes, combined or separated, to precipitate soluble toxic elements, rendering them relatively inert, thereby reducing them to their solubility limits in the liquids accompanying process solid wastes. Dispose of process solid and liquid waste directly, without removing soluble innocuous compounds or remaining traces of toxic elements; • Precipitate toxic elements, thereby stabilizing them as solid waste, wash process solid wastes free of soluble innocuous compounds and remove the remaining trace amounts of soluble toxic elements. Dispose of the washed solid wastes as desired, as well as of the recovered, concentrated (solid) innocuous compounds and toxic element streams as additional process wastes . • Treat process wastes as above, except that the bulk solids are dried prior to disposal. As may be seen in Table 11.1, toxic elements are widely dispersed in the process waste streams. Furthermore, the largest volume waste streams in all processes, leached tails, slags and lime-boil solids, all contain toxic elements and, therefore, must be treated for any disposal option other than containment in lined tailings ponds. Thus, there is little to be gained by separating process wastes prior to treatment and this option has not been considered further. The majority of the toxic elements leave the process fixed in solid process waste streams, exiting mainly in the solid residues from leaching, precipitated, and fixed in neutralization and lime-boil steps, and incorporated in slags. The disposition of each element in the various processes is largely inherent in the process chemistry, and very little can be done to alter its ultimate fate without major structural changes within the process. Some changes can be made on the distribution of toxic elements which had been solubilized. However, more effi- cient washing of leached tails would tend to retain toxic elements within the process loop where they would be removed, to a certain degree, in subsequent precipitation steps. No sequence of washing steps is completely effective, and 11-5 attempts to reduce concentrations by even one order of magnitude would be extremely costly. Furthermore, some purge point must be provided within all the processes for the elimination of soluble innocuous materials, such as sea salts, to prevent their build-up to saturation levels. Such an occurrence would cause significant operating difficulties, since their precipitation could not be, in general, controlled and would adversely impact process viability. Thus, it is not likely that significant changes could be made in the nature of the process waste without major modifications of the process routes; and since the waste products would still require further treatment, major process alterations were not considered further as basic treatment techniques, per se. The waste treatment alternatives considered further include precipitation of soluble toxic elements with direct discharge of treated wastes, precipitation of soluble toxic elements followed by washing of solid waste to remove soluble innocuous and toxic elements for separate discharge, and precipitation of toxic elements followed by washing and drying of solid wastes for discharge. Preci- pitation of toxic elements can be achieved by treating the combined process waste streams with slaked lime. Sufficient lime would be added to remove soluble sulfate as gypsum and raise the solution pH to a sufficiently high level to precipitate metal hydroxides. It is assumed that, under these condi- tions, anion-forming elements such as arsenic, vanadium, and molybdenum would also be removed, although this is somewhat more problematical. Other metal hydroxides, such as magnesium, would also be precipitated, but alkali metals and alkaline earths would remain in solution as chlorides, hydroxides, and sulfates to their solubility limits. An excess of lime can be added to stabilize the precipitated toxic elements, including heavy metals, against redissolution on contact with acid rains during subsequent disposal. It is possible, however, that additional treatment would be required at the disposal site to ensure that these materials would not be resolubilized on standing for long periods of time in contact with possible chemical environments which cannot be predicted at present. For the option requiring removal of soluble, non-toxic compounds from the processed solid wastes, washing by counter current decantation in a thickener train similar to that used in tailings washing can be used. Washing would be carried out with reclaimed, essentially solids-free water, so that the liquid in the washed wastes slurry would contain on the order of 250 ppm innocuous dissolved solids, which is approximately the level assumed for the process make-up water, and of the order of 1 ppm total content of toxic elements, including heavy metals. Soluble toxic elements would be removed from the wash water by sorption on activated carbon or a suitable ion exchanger, which can be regenerated by hydrochloric acid. Recovery can be made by evaporation and crystallization for disposal as solid chloride salts. Salt-free water for recycle would be recovered from the dilute solution of soluble innocuous salts by reverse osmosis and evaporation with the innocuous salts rejected as solid chlorides and sulfates. Solid waste drying, if required, can most economically be achieved by direct contact with hot combustion gases. 11-6 11.4 PROCESS DESCRIPTION The generalized waste treatment process considered to be "most likely" is shown schematically in Figure 11.1. Detailed material and energy balances are developed for each of the five treatment processes considered in this study. These are presented in Figures 11.2 through 11.6 It is important to note that in the very process of further treating the wastes for the removal of soluble toxic and innocuous compounds and for waste- drying, additional process wastes are created and energy consumed. Lime is added to precipitate toxic elements and soluble sulfate, forming gypsum. Also, additional coal must be burned to supply the required energy for water evapor- ation, which, in turn, generates not only more combustion ash, but stack-gas scrubber solids. Lime is slaked with recovered wash water, or in the relatively clean tower and boiler blow-down streams, and mixed with process solid wastes. Sufficient agitation is provided to mix the slaked lime with the waste slurry, and residence time is provided to allow the precipitation reactions to go to completion. It is anticipated that large retention volumes will be required in view of the difficulty of removing the toxic elements to very low levels. The process solid wastes, now including precipitated toxic elements, pass to a solids washing step where soluble constituents are removed by counter-current washing with reclaimed solids-free water. The wash water, containing soluble innocuous and toxic elements, passes to a heavy metals removal step where the majority of the remain- ing soluble heavy metals and toxic elements are removed by sorption on activated carbon or an appropriate solid ion exchange material. The sorbent is regenerated and the heavy metals are removed by contacting the sorbent with aqueous hydro- chloric acid. The solubilized toxic elements, in diluted acid chloride solution, pass to an evaporator where water and excess acid are recovered and the toxic elements are precipitated as their chloride salts. Made-up and recovered hydrochloric acids are recycled to regenerate the sorbent. Spent sorbent is sent to the main boilers for combustion and the removed heavy metal and toxic element chloride salts are treated as required and sent to disposal. The raffinate from heavy metal removal are solutions containing dissolved salts at concentrations between 0.3 and 3 weight percent. These pass to a reverse osmosis step. Water is removed with a 97% salt rejection to produce a concen- trated brine of approximately 3% solids. The brine, in turn, passes to a multiple-effect evaporator crystallizer , employing between four and eight stages, where water is removed and the soluble metal chloride and sulfate salts are precipitated. The recovered soluble salts are treated, as required, for disposal. The condensates produced in the evaporator crystallizer are combined with the desalinated water obtained in the reverse osmosis step and recycled to wash the solid wastes free of soluble salt in the CCD circuit. Flocculent is added, as required, to obtain the desired settling characteristic. The process solid wastes, washed free of soluble innocuous and toxic elements, pass to a dryer where water is removed by direct contact with hot combustion gases. Off-gases from the dryer pass through dust removal devices and to the main stacks for treatment and for removal of sulfurous compounds. The dried solids, still containing sufficient moisture to prevent excessive dusting, are conveyed from the dryer to disposal. Additional facilities pre required in the materials handling area to receive the extra coal, lime, 11-7 £ -• ^ ^ 5 ,^ i^ ^ ^ ^ U ^ iu !-0 kt o O CO CO CO UJ o o CL Q UJ M _J < UJ Ld Ld in q: LU f- co < r' ^1. ■ .^ — — IT J--.. f - S J- ■t' K ; \ I W <, .^ V '^O' •«- «.. - l-i . /"L.. ■^ i S^- U -J Hi •4! 11-9 I ! s ~^r >o --1 >s f\ll >0 — 09 ^- Oi f\l Oj "»7 S3 ^ iQ oj ?: <^ W, c^ o ^ •^ ^ ^ J 11-10 11-11 ^5-.* ^"^ ^ 1. 1 " i; i i >iil-: \'- T - — H V-^- I — ' : y ,'^ 1-£:1 ■t>, jh^l p 11-12 11-13 f locculant , and cooling tower and boiler treatment compounds required for use in waste treatment. Facilities are provided in the service area for circulating additional cooling water and raising the additional steam required in the evaporation step and for combusting the additional coal required to generate hot gas for solids drying. Additional stack gas treatment equipment would be required as well as additional process power which would be partially off-set by power raised internally in generating additional steam. 11.5 IMPACT OF WASTE TREATMENT ON PROCESS MATERIAL AND ENERGY BALANCES The materials and energy requirements and effluent production summaries for the three waste treatment alternatives considered are summarized in Tables 11.3 through 11.7 for the five nodule treatment processes considered in this study. In all cases the simplest alternative, limlng-out soluble toxic elements to fix them as inert precipitated solids, involves only negligible changes to the overall process material and energy balances. Such an option could easily be considered in a minor modification of the base case production scheme. Its incorporation would require only additional slaking facilities and perhaps a larger waste containment surge vessel on the plant site. The objections to defining waste treatment as simply consisting of this procedure, however, are two-fold. First, the process liquid waste will still contain small, though significant, amounts of soluble toxic elements. Second, these soluble toxic elements, as well as soluble innocuous compounds, could still be leached and these materials introduced into ground water systems if not impounded in lined containment facilities. Thus, this scheme does not significantly affect the problem of safe disposal of process wastes. The waste treatment sequence which involves washing the solid wastes for removal of soluble innocuous and toxic elements imposes modest increases on the material and energy requirements for the overall process and would require significant additional capital expenditures. Significant amounts of additional by-products (wastes) are also produced by this option, notably the recovered toxic-element metal chlorides and the precipitated soluble salts produced in the evaporator crystallizer . While the former wastes have been classified as toxic, their chloride salts are not extremely hazardous to handle and the pro- duction rate is relatively small. These materials could be further treated to render them less mobile, by encapsulating them in a polymer matrix, for example, to render them more suitable for disposal. Disposal of the recovered soluble salts on land would require the use of a lined containment area to prevent their leaching into ground waters, but the required lined containment area would be relatively small. However, since these materials are innocuous, being mostly sea salt in some cases, it is likely that the disposal at sea would be a viable alternative. While disposal of the washed solid slurry may no longer require lined containment ponds, the area requirements for land-based disposal would be essentially the same. 11-14 •H X "* W o « 13 H 7^ •H ^ 1— 1 a O i-j rs cyj ri tfi 4-1 4-) > •H r^ u o. g a •H E U a Xl (U rH c J-i W C3 Ph ro 00 r- r^ . 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O CX a. iH cn'^ 4-1 CO rH 4-1 4-1 *. rH CO C c CO CO GJ QJ QJ 01 OJ 4J 4-1 AJ rH d c cn rH XI o o CO CO D •\ o CJ 3 4-1 I— 1 cn 0) o o 4J cn T3 s; v3 4-1 rH rH •H cn CO CO rH >% X) CO CO XJ O > -H S QJ CO CO rH CJ 2 a) o -3 f-H 4-1 33 CO •H J3 >^ !-i 3 3 > Ql >. >> Cr rH C3 C u u .-^ o QJ M o o -^ CO X 11-15 CO H en Cfl QJ a o S-i p-i X CJ CO cu hJ -o •H >. 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CO CG CO :s )-i )-i c c U 4-1 60^ >. >. >. ^~, --^ 4-1 4-1 -^rn ^ H ,H 4-1 en ■n ro - 4-1 QJ ■^ C iH r-( C - B-S CO 4-1 r-~- rH ■U (3 CO a r« *. - •> OT QJ ^-^ M U rH QJ QJ ^ QJ QJ CO e e >-i rH ti ^ -H -H rH CO UrJiJCOSfn rSpL, 11-16 CO m CN m CM in o^ rH rH 00 si- rH '-^ C 4-1 •H cn B 4-1 O u 3 " >% Vi C ■--, >, O CO ~^ a -H (3 rH W 4-1 O CO cn CO 4-1 &0 u 1 Oon ^ O P-O o H CO rH rH > tn V4 cn 0) QJ CO QJ CJ H cn CO CO :3 13 00 12 QJ (3 ^ -H 13 O rH -H CO O rH AJ O O -H CO C_) CO kJ 13 •H d cr M >> >^^ cn M W C ?^ >. 4-J cn cn C Cc-o VI >^ 4-1 4-1 rH ■ ^^ ,-\ e B CO ** CO a. cu cn 00 ex a, rH •^ 4-1 cn r^^O ^ »» 4-1 CO 4J 4-1 •N r-{ COrH C c cn CO QJ 0) QJ CO QJ 4J 4-) 4-1 rH e c cn rH ^ CO CO 3 *> u CJ 3 4-1 rH cn QJ 4-1 cn 13 ^ CO 4-1 ^ rH •H cn CO CO rH >. 13 ^ CO 4-1 > •H s CO CO rH QJ S QJ -0 rH 4-1 K CO •H ^ >. 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C/'J u !-i c >. >-. 4J en en C cm J-l >^ 4-1 4-1 I-H ■^^ e S ^ iH a a< m en CO a Q. „ 4-1 tJC |-^ en 1—1 •s «« 4-1 CO X) 4J 4-1 ^ i-H CO c c en CO 1—1 cu OJ cu en cu u 4J 4-1 , — 1 c c en i-H ^ CO CO #« u 3 4-) iH en CU 4J en -a S to 4-1 r-l r-l •H en CO CO .H >^'a rj cn u > •H 3 01 CO CO r-l a S cu ■a 1—1 4-1 3:; cr. -^ X:. >^ M D d > CU >> >s C" iH CO c u )-! — * CU M Q C — ^ C/5 ■X 11-17 •H X f CC O co 11 H CO •H ^^ I— 1 O O 4-1 #^ C/3 CO en 4-1 4-1 >, •H C S-i CX OJ o •H g tJ en CM o 00 rOfnroOCNOLn~3- m in vO in £N o in ro o <3^ ^ en CO o g o 3 C^ •H QJ ^ e •H W 4-1 4-1 -a c CO e 4-1 en CO CO u •H H (U a) 4-1 en CO >a •H en X CO o :s H QJ d +J CO to CO TJ +j en -H •H -W r-l o. G o •H QJ C/3 cj 6 QJ QJ P-i W CJ >< E^ o i-H H c o Q) 4-1 en CO u 4-1 C •H QJ Dh B •rH QJ o rH Q) W >-i CM SO rO C?^ vD '"^ t^ r-~ ro CNI ^ O rHrnOOCNiOO-^ Csl iH r-^ in ^-^ CN » in cNi CN C QJ e QJ U •H Si QJ Pi 4-1 c CO QJ U •H 5l en a ro O >-i !-i >-i >^ >. >. ^■^.^ eo tn tn c u u u C (H O !>^ >. o o >, 4-1 ---. !-l ^^ 4-1 4-1 -~^ ,-{ s: >, en enoo CO 13 ^ C n en C O 00 1^ ec o o o O r-\ O 4J vp VO 4J 00 ■^ 4-1 C S~5 CO r~~ rH OO 3 " « ^ » en QJ •—' u u yt rH QJ QJ ^ O Q) QJ CO e 6 M rH O 4-1 £: O -H -H O U rH CO O CJ H-5 K-) C/2 s: tn s PL, CN OO O en QJ en CO O 0) CJ CO OO o u ^ >^ ^ g tn --~- 2 C rH •;J o CO tf 4J ClO CO goo g-rH rH CO > j-i - en QJ en QJ 5 QJ -u O 4-1 en H en CO to 3; 60 s C •H 13 rH -H O r-^ o o o in in o ^ Vh tn >^ V4 o eo >. 4-1 c ■ o 4-1 en CO C o O rH ) 4-1 rH 6 B o " CO D. Cu, l-i en tjO Cu a " 4-J eo rH»0 •« ff^ 4-1 CO O w 4-1 #^ rH cn ,-H C C en cO G (U QJ in QJ 4-1 4J 4-1 rH c c CO rH J2 o o CO CO 3 *• u o 5 ■P rH cn 0) O CJ 4J en x) S to 4-1 rH rH •H en CO CO rH >.13 CO CO 4J o > -H 5 a c/: CO rH QJ S QJ O 13 ^ 4-1 !i: C/3 •H XI f^ u 3 3 > QJ >^ >v cr ^ CO c )-i U •H O QJ M o o ►J e/3 K 11-18 LO vx> r- r-H o cx) c^ o rH 00 '.O in H tH O tN m u-i (J\ iH ro tn o cTi ro fo 00 tN 1,0 0> ^ vD O vD LTi -i 0) J-l CO 3 u 13 "• — C 4J cn •H m C g 5-1 5-1 O ~->ro >^ >^ 4-1 4J o '^ — ^^ IH iH 5-1 CO C COiO C o 5-1 3 >, 5-4 o O rH ■^^ U C tN 4J 4J ^ e e J_l o tn ^ CO (X a ^* 5-1 >^ ■r) 13 -H C rH m ^. 00 a a >^ >^ . >. o o >. -p ;?. o O D.O o ID tn CO OJ (1) ~~^ ■--^ -U -U i^ >H CO rH .H O 0) C/D OJ ;-> 4-1 en w coco .H ^ (^ > a 4-) rH C C p: Co-) fo C O CO S w 3 cn rH 43 o o o o o o O rH 00 yl 0) #% ^ -a CO CO 3 ^ o o ■W 4J iH >H 4J QJ m }-i *\ U) o 3 4J iH cn »o ^0 4-1 0) 0) cn Qi 5-1 (1) O CJ 4-1 (0 ro CO CO o o cn tn s (U 4.) p-i X) X "^^ 4-J rH rH O O « " U rH iH CO CO o 4-1 CO •H cn CO cd rH rH 0) ^-^ c 5 a H U) CO cn ,-H >>13 CO c/: 4-1 C B-5 CO CO 3 OJ O > -H ^ n 3 " - CO > rH 0) OJ ^ O QJ OJ O O rH •H 3 IS 5-1 3 3 > CO e 6 M rH O J-" 3 •H CO O iH cr QJ >s >^ O- rH CO O -H -H O u rH n) o 4J 4-1 O O •H rH C 5-1 U •^ o _CJ CJ ^J ^J C/3 3 tH 12 CL- •H < tn cj LO hJ a 4-1 o H M O C C/2 X 11-19 It should be noted that the recovered soluble salts would be similar in nature to the much larger amount of fused salt solid wastes produced in the reduction/hydrochloric acid leach process. These salts, however, contain significant amounts of toxic elements, particularly barium and lanthanum, which amount to almost 2% of their total weight. Thus, disposal of this waste material presents a unique problem. The volume is larger than can be polymer encapsulated conveniently. However, it could be easily contained in lined impoundment areas. Since the material is fused and will have been cast into relatively large-sized blocks, it will not dissolve rapidly. One pcssibJe: treati.-crt alternative, though expensive, would :.nvolve size reduction and redissolution of the chlorides in the process liquid waste stream. The toxic elements would be removed by precipitation with lime and the soluble chloride salts re-recovered by reverse osmosis — evaporation/crystallization. In this case, the salt should be suitable for disposal at sea. In view of the lack of operating data, the nature of the assumptions re- quired to estimate the concentrations of soluble toxic elements and the uncer- tainties in the process chemistry of these non-ideal solutions, the waste treatment scheme proposed must be considered conceptual in nature. It may be very difficult to obtain the required degree of mixing of slaked lime and waste slurry to ensure complete precipitation. Reduction of heavy metal and toxic elements below their solubility limits by carbon or ion exchange sorbents would represent an extension of this technology to a much larger scale than is presently utilized. Also, fouling and scaling problems in the evaporator/ crystallizer could present severe operational problems. The energy requirements imposed in removing water from washed solids are significant and would probably represent a major economic penalty for most process options. Furthermore, since the dried solids would themselves present disposal problems not encountered with a tailings slurry, such as dusting, it is not clear that this waste treatment option would ever be exercised. Implementation of either the solubles removal or solubles removal/solids drying treatment schemes would require an additional 20 to 30 acres of land at the plant site to accommodate the washing thickeners and other equipment. Some additional operating personnel, approximately 20, would also be required, along with facilities for disposing of the additional wastes created. 11-20 12.0 NOISE ASSESSMENT 12.1 INTRODUCTION Five possible chemical processing schemes for the recovery of manganese and other metals from ocean nodules were reviewed for their noise emission characteristics. These processes are identified as: 1) Reduction/Ammonia Leach Process; 2) High Temperature Sulfuric Acid Leach Process; 3) Reduction/Hydrochloric Acid Leach Process; 4) Smelting Process; and 5) Cuprion/Ammoniacal Leach Process. Detailed designs for the nodules processing plants have not been developed; hence, details of equipment types and usages are not available. Thus, noise levels from these processing plants can only be estimated. The major noise processing equipment can be grouped into the following general categories: 1) Fluid handling equipment (i.e., pumps, mixers, agitators, etc.); 2) Gas handling equipment (i.e., compressors and fans, etc.); 3) Material handling equipment (i.e., mills and conveyors, etc.); and 4) Energy conversion equipment (i.e., combustors) . The power requirements of major equipment categories are estimated from designed processing rates for various plant unit operations. These data were obtained directly from the detailed flow sheets and line tables. Based on the power requirements of the equipment, estimates of their noise levels are made using data from published literature (Harris, 1957, and Magrab, 1975) and Dames & Moore files. However, it should be noted that since equipment types cannot be specifically defined, noise levels are only "order of magnitude" estimates. 12.2 NOISE SOURCES Equipment emitting high noise levels is either enclosed in a building or located outdoors. In this analysis it is assumed that fans, compressors and pumps will be enclosed by housings with approximately 30 dB attenuation. Induced draft fans and cooling towers will also be quieted by silencers with approximately the same attenuation capabilities. Indoor noise sources will be subjected to an additional attenuation of approximately 20 dB resulting from an assumed metal plant building. It is, therefore, anticipated that indoor noise sources will not contribute significantly to the outdoor ambient sounds. The major noise-producing equipment types and estimated noise levels for the five processes analyzed are presented in Figures 12.1 through 12.5. The principal noise sources contributing to the ambient sound are unenclosed conveyors, feeders and fans. 12-1 o n S 4J 0) Q) fa O o o H o in -P 0) > •d c o CO -o Q) +J X! &^ •H (U I < o o CO a< g 3 • (l4 M (U -p fd (0 IS to u H U 0) iH !T« o ^ •H C >i o s: •H > PQ >i m C M >^ rH to (p i IS -P rH fd (d c c CO (U (U o iH u (t3 m ffi (U c fO C M to 5^ -P rH CO CQ o EH & S^ •H fO cu u a (0 PQ $-1 > to rH EH-H ■p >1fa e e o »^ o CO tn CO (d (d +j (0 M >1 O CO >1 O o ^1 C to cu •H to m s-< (0 M o 0) U CQ fO to (U 4-) CO (U (0 (D •rH QJ Sh (d r, (0 g (U > c tn a > fC a d:; tt 15 rH to O rH 0) Cnx) a ■3 >i S-l c J-l W M-l E c; tn s-l to c (d 4-) 14H g (D >^ u ■H •H •H (d m :3 C QJ 3 fO 3 <-i CD fO U (d 14H o u 0^ Q < U +J <: o o PM u -H <; Oh -H e^ m U U fa Ph CJ S O CJ cu o u > QJ -o (0 u (-1 (U QJ 2 OJ O (xi i-:i <3 CO w cj cc; •H fa 12-2 ,— » o < VD m S +j •^ 0) H^l •0 c :3 w o Tl ro 0) +J j:: Di •H 0) 1 <: o CN rH tn c iH a &» •H e ■H c s :3 ■P -rH 0) w a, (C -P Q. >i Dj N fO E U B T! •H rH d (0 3 H >H S: TJ tr> fd Cn 0) >i 5 S CHT! cn m c c c ^ C 4J u O C OJ C u Ch (C -H 0) -p •H fd QJ Eh Eh •rl 0) m cu e •H -P U X 4J -C c > +J fo t, 3 3 -p >i (0 0) C Cfi -rH O CO tji tji n3 CU (0 (-1 M >i (1) tfl IW u u C C rH Vh cn rH u (Odd) ^ B V) S cn M-i cn Q) in (u •H -H D 0) m PQ • a e u > o nH Ou a, (fl a, 0^ a. s rH rH U rH CP u a, •H H C tr> (d w -f e cn g 0^ B 6 tn O M -H UH s-i cn (1) (0 •rH a £ M h (U ffi nj g (D o h^ a, tH u <: cn (0 04 04 4-> C 0) e a ■H TJ d T3 ■H CT C u W X! P 1 CO CD d) o !h U) H-) 3 •H H-) cn O cn CO cn 2 C Sh (U ■rl (U o Xi H-) a (D 3 E >H H-) X! CD 04 (d ■H Eh e M 1 x: •H H-l x: u -P c cn rcj C/5 •H d) W U E h^ CN CN ^^ -a c O w CD -P Cn •H (U 12 I < c o in o o O o •H -P U a^ QJ o iH iH ■H s >1 u c o u 0) cn c E •H 0) o o U U (U 0) u o cn (L) U Cu e o u u Q < fa M C C -H fd fo <; fa fa w cn c Cn cnM 4-j m O 1+-J M-i O O O U C o QJ 1-:) cn a, Oj 0) > o o cu Pi 0) tn 0) c re tn S-l o >1 tn Q) a > e c c 13 O fO 0^ CJ fa >1 QJ > o CJ QJ QJ QJ G C > > S^ O O O 3 -rH U U -P +J W QJ QJ QJ Oj S-l Di K (15 5-1 QJ O 5 r-^ CN CM ^ O ^ ^ ^ XJ -H u CJ o <; oQ 5-1 M o CJ O tn cn QJ s-i tn & Ch E E O d S O Q iH ffi O fii QJ i-i U Eh O >i tn QJ fd > iji C 4-1 o u-i O O B u OJ -P c fd fd :2 fa Cn Cn G C •H -H ■P -P (d fd d 13 u u Sh u ■H -H CJ u u u QJ QJ ^ IS o o Cn Cn C C ■H -H o o o o u u QJ a o U X! O E M O ft O E ft S-l QJ QJ O QJ nH fa CQ U U QJ QJ r-^ r-\ •H -rH o o c 3 O CO -p c; 0) ■H O -P Cn ■H 4-> 13 St •H Vh •p o u -P tn G tn QJ QJ E O Cu o ■H !h 13 ft Cr' W T! ■H O < tn u -H -H QJ M > O QJ rH U QJ O cn !h •H x! o >, 2: X \ TS C 0) O 4-1 -H fd +J E U •H 3 -Pt3 tn 0) W (X CM QJ U P Cn •H fa 12-4 o -p 0) [n o o o (0 0) > hi o in •o a) -p •rH 12 o IT) o o en o •H s 0) >1 >H 5 U M CU m tfi W 0) ^ rH T3 CI, w Qj 5 pa C e C B r-^ O 3 m 3 rH m -p U cu Cn p^ PQ c; (U >H QJ CO U U U V) -r^ B 0) QJ QJ (T3 < -P n3 >i •H ^ ^ -P CJ (C5 C C M u 4J iH (fl C QJ mow W QJ d T! tfl Eh EH S C O Vj -H CO >^ M > 0) C M -H H -p >, V) g (CJ (ji Cri w -H -P 03 U (D >> M >i u -p ffi >1 C c w P w en U (^ U Q) QJ (1) w QJ CO tfl en 0) -H -H QJ U 3 (d rfl 6 a, > C ^ > Q. cc; a, Qj cn a, > rH rH U 3 -Q CP a ■H g c o c tP e S ^ C C C C rH e c en O T3 E ^ 0) MOO •H rH C d (d 3 H fC (C (d 13 td d QJ >H Q) 4-1 M O) u u +J PQ u -rH a, •rH cu pLl pL| Pl, (i( -rH CU u o U U pL, « U O CU U 3 O p 0) •H > 0) T) fO p U) +> ^^ M (U 0) P nJ (0 (U O « ^^ < & s w c QJ e a -H P Ti tJ' C W 4-1 CO -p en iH rH QJ Q) ■H > X! QJ 1 cn QJ en cn -P QJ ■H U Cr>0 2 C iH •H fl, T3 -P Q) d cn -P X5 c fO •H •H g H H-) •H 4-1 rH +J C QJ W g H U CO ^^ «* >Cvl iH QJ U d tn PLI 12-5 CQ o -P 0) O o o o in 0) > 0) Ti O 03 TJ CU +J x: ■H OJ I <; o o 00 O W H U U o •H •P CU M (U u o M •H s m >i a i^ E 03 a Xi d, C n o i m w SM rH >1 ^ rO 13 0) 0) g u > s ■H H c U -rH r-H eu u u CQ en w m to o 0) w CU Cli h en tp C C -H ■H x; X3 u U fO ft! CU OJ J 1^ \ CO c 0. C o E •H 3 -H ■p Oi -P o rO 3 n Ti •H CD X Pi o CM u (D -P 03 [2 W U) (U o o u u o m (U M U fl, a e o u > a c e o ^ en tP fO fO o S [2 CnH Cn tn C C C -H CO H -H O fC H -H Id tn O 'O M-l O (U IH O^ U U cc: O m fc fO fC •H -H OJ 0) +J -p s s 12-6 12.3 NOISE LEVELS The noise levels generated during the operation of each processing plant is estimated by combining, on an energy basis, the sound energy of each individual noise source. The range of A-weighted sound levels at 100 feet from the center of the plant noise sources, for each process area, are presented in Table 12.1. The noise level is estimated to be highest for the reduction/ammonia leach process, where the noise level is approximately 56 to 66 dBA at 1000 feet from the center of noise sources. The high temperature sulfuric acid leach process is estimated to contribute the least to ambient sound. It's noise level is estimated to be 48 to 58 dBA. Preliminary evaluation indicates that process- ing plants will be rectangular in shape on sites approximately 2400 to 3200 feet in size. The noise levels presented in Table 12.1 can conservatively be interpreted as the noise contribution from the processing activities to the ambient sound at the plant boundary. The A-weighted sound levels of typical environmental noise sources is presented in Figure 12.6 for comparison with anticipated noises produced by the five processes as presented in Figures 12.1 through 12.5. Table 12.1 Estimated Noise Level At 1000 Feet From Center of Processing Activity Range of A-Weighted Sound Levels (dBA) at 1000 Feet From Center of Activity Reduction/Ammoniacal Leach Process 56 - 66 High Temperature Sulfuric Acid Leach Process 48 - 58 Cuprion/Ammoniacal Leach Process 51 - 61 Smelting Process 52 - 62 Reduction/Hydrochloric Acid Leach Process 54 - 64 12.4 REFERENCES CITED Harris, C, Handbook of Noise Control , McGraw-Hill Book Co., 1957 Magrab, E. , Environmental Noise Control , Wiley Interscience , 1975. 12-7 TYPICAL A-WEIGHTED SOUND LEVELS At a Given Distance From Noise Source Environmental Decibels RE: 20yN/m 140 50 HP SIREN (100 FT) JET TAKEOFF (200 FT) 130 120 RIVETING MACHINE HO CUTOFF SAW I PNEUMATIC PEEN HAMMER TEXTILE WEAVING PLANT SUBWAY TRAIN (20 FT) PNEUMATIC DRILL (50 FT) FREIGHT TRAIN (100 FT) VACUUM CLEANER (10 FT) SPEECH (1 FT) LARGE TRANSFORMER (200 FT) SOFT WHISPER (5 FT) 100 90 80 70 60 i I 50 40 30 CASTING SHAKEOUT AREA ELECTRIC FURNACE AREA BOILER ROOM PRINTING PRESS PLANT TABULATING ROOM INSIDE SPORT CAR (50 MPH) NEAR FREEWAY (AUTO TRAFFIC) LARGE STORE ACCOUNTING OFFICE PRIVATE BUSINESS OFFICE LIGHT TRAFFIC (100 FT) AVERAGE RESIDENCE MIN LEVELS RESIDENTIAL AREAS IN CHICAGO AT NIGHT STUDIO (SPEECH) THRESHOLD OF HEARING YOUTHS 1000-4000 Hz 20 10 {0 STUDIO FOR SOUND PICTURES Figure 12.6: Typical A-Weighted Sound Level As Measured by a Sound Level Meter 12-8 13.0 GLOSSARY Agglomeration - In ore benef iciation, a concentration process based on the adhesion of pulp particles. .Loosely bonded associations of particles and bubbles are formed which are heavier than water; gravity is used to separate the agglomerates from non-agglomerated particles. Ammoniacal - Of, relating to, containing, or having the properties of ammonia. Anion - A negatively charged ion, such as a hydroxide, a chloride, or a sul- fate ion; opposite of cation. The ion in an electrolyzed solution that migrates to the anode where it may be discharged and liberated. Anode - The electropositive pole. The electrode at which electrons leave a device to enter the external curcuit; opposite of cathode. Aqueous - Of, relating to, or having the characteristics of water; watery. Autoclave - A closed vessel for conducting chemical reactions under high pressure. To treat in such a vessel. Briquetting - To form loose, powdery substances into a compact form (briquette), usually by compression with or without the use of heat or other agents. Calcine - To expel volatile matter by heat, as carbon dioxide, water, or sulfur, with or without oxidation; to roast; to burn. Carbonation - The process of introducing carbon dioxide into a solid or fluid. Cathode - The electrode where electrons enter (current leaves) an operating system, opposite of anode. In a battery or electrolytic cell, it is the elec- trode where reduction occurs. The product formed by deposition of metal by electrolytic reduction. Cation - A positively charged ion which migrates to the cathode or negative terminal in an electrolytic cell. An ion having a positive charge, such as cupric, sodium, or hydrogen. Cementation - The removal of an ion in solution by reduction with another metal lower in the electromotive series, as the reduction of cupric ion from solution with simultaneous disolution of metallic iron. Also called displacement. Centrifuge - A rotating device for separating liquids of different specific gravities or for separating suspended particles, such as clay particles in an aqueous suspension, according to particle-size fractions, by centrifugal force. Chelating Agent - A substance which contains two or more electron donor groups and which will combine with a metal ion so that one or more rings are formed. In metal finishing, a chelating agent is used to control or eliminate certain metallic ions present in undesirable quantities. In extractive metallurgy it is used to remove selectively one ion from a mixture in solution. 13-1 Clarif ier - A centrifuge, setting tank, or other device for separating suspended solid matter from a liquid to produce an essentially solid free liquid stream and a more concentrated solids stream. Classifier (air, rake) - The term classifier is used in particular where an up- ward current of air is used to remove fine particles from coarser material. The rake classifier is a device that takes ball-mill discharge and separates it into two portions by mechanical means. Coextract - To withdraw together, not necessarily in equal portions, substances (as dissolved metals) by physical or chemical process; also, to treat with a solvent so as to remove soluble substances. Condenser -Absorber - An apparatus for removing heat from a gas (steam) so as to cause the gas to revert to the liquid state to be absorbed by more liquid such as water. Countercurrent Decantation (CCD) - The clarification of wash water and the concentration of tailings by the use of several thickeners in series. The water flows in the opposite direction from the solids. The final products are slurry which is removed as fluid mud and clear water which is reused in the circuit or, containing valuable constituents, is sent to further processing. Cyclone - Refers to the conical-shaped apparatus used in dust collecting opera- tions and fine grinding applications. In principle, the cyclone varies the speed of a fluid which determines whether a given particle will be removed through force of gravity or be carried through friction. Demisting - The removal of liquid droplets suspended in a gas. Densif ication - To increase the density or concentration of a substance. Dissolution - The act or process of dissolving or breaking up; as a separation into component parts, usually into an aqueous solution. Electrolyte - A nonmetallic electric conductor (as a solution, liquid, or fused solid) in which current is carried by the movement of ions instead of electrons with the liberation of matter at the electrodes. Electrolytic - Of or relating to electrolysis or an electrolyte; produced by electrolysis. Electrostatic Precipitator (ESP) - A device for removing dust particles from air or gases. A high voltage is maintained between a grid of elements through which the dust laden gas passes. Ionization of the gas permits charged dust particles to migrate to collection points and be removed. Electrowinning - Reduction of metal from a solution by means of electrochemical processes. When sufficient voltage is applied, metallic ions in solution are reduced and deposited at the cathode. 13-2 Endothermic - Accompanied by the absorption of heat, opposite of exothermic. Evapora tor /Crys tall izer - A specially designed apparatus that crystallizes materials in solution. Supersaturation is produced by evaporation, usually without appreciable cooling. Exothermic - Characterized by or formed with the evolution of heat, opposite of endothermic. Fayalite - A silicate of iron, Fe2Si04, belong to the chrysolite group, Flocculant - An agent that induces or promotes f locculation, or produces floccules or other aggregate formation. Flotation - The method of mineral separation in which a froth created in water by a variety of reagents floats some finely crushed minerals (typically surface active sulfides) whereas other minerals sink. If the particles are retained in a foamy layer several inches thick, the process is called froth flotation. Air flotation is carried out in a vessel agitated by air. Fluxing - Fusion or melting of a substance as a result of chemical action which causes the development of the liquid phase by the admixture of low melting components. Fugitives - Light particles of material treated in roasters, smelting furnaces, and converters that are inadvertently carried out of the unit suspended in the off-gas. Fused Salts - Salts,, that is, ionic compounds, in the molten state. High tem- peratures are usually involved in maintaining the molten state. Gangue - Undesired minerals associated with ore, mostly nonmetallic. The non- metalliferous or nonvaluable metalliferous minerals in the ore. Gypsum - A hydrated calcium sulfate, CaSO^ -21120. Hydr ochlor ina t ion - The process of treating a material with hydrogen chloride to produce a physical or chemical change or both. Hydrocyclone - A cyclone separator in which water is used. Hydrolysis - A chemical reaction involving alterations of a substance with the addition of the elements of water. Hydrometallurgical - Pertaining to hydrometallurgy , the treatment of ores, con- centrates, and other metal-bearing materials by wet processes, usually involving the solution of some component, and its subsequent recovery from the solution. 13-3 Ion Exchange - The exchange of ions contained in a solid for different ions in solution without destruction of crystal structure or disturbance of electrical neutrality. The process is accomplished by diffusion and occurs typically in solids possessing one or two dimensional channel ways where ions are relatively weakly bonded. Ions may also be exchanged between two liquid phases, as an aqueous and immiscible organic phase. Laterite - Red residual soil developed in humid, tropical, and subtropical regions of good drainage. It is leached of silica and contains concentrations particularly of iron oxides and hydroxides and aluminum hydroxides, and is an important source of nickel. Leaching - Extracting a soluble metallic compound from an ore by selectively dissolving it in a suitable solvent, such as water, sulfuric acid, hydrochloric acid, etc. The solvent is usually recovered by precipitation of the metal or by other methods. Lime Boil - A chemical reaction between lime and ammonium sulfate where steam is introduced to the reacting slurry. Ammonia is released and gypsum is formed. Liquor - A liquid substance, usually an aqueous solution containing dissolved metals or salts. Lixiviant - A reagent to extract a soluble constituent from a solid mixture by washing or leaching. Matte - A metallic sulfide mixture made by melting the roasted product in smelting sulfide ores of copper, lead, nickel, and zinc. Mill - In mineral processing, one machine, or a group, used in crushing and grinding, i.e., comminution of ore. Mixer-Settler - A combination of a mixing and settling apparatus into a single unit or stage and the combining of a number of stages to create a process that allows for interaction by mixing and separation by settling to produce a more complete or thorough reaction between aqueous solutions and organic liquids. Organic - Being, containing, or relating to carbon compounds, especially in which hydrogen is attached to carbon whether derived from living organisms or not. Overflow- En trainment - The particles remaining in a solution that is taken from the top of a settled solution. Oxidation - Combination with oxygen; increase in oxygen content of a compound; increase in valency of electropositive part of compound, or decrease in valency of electronegative part. A reaction in which there is an increase in valence resulting from a loss of electrons. Polish Filtration - A refined filtration of a fluid to a very low level of sus- pended particles. Precipitate - To separate a solid form from a solution by chemical means; the newly formed, soluble solid so produced. 13-4 Pregnant Liquor - A value-bearing solution in a hydrometallurgical operation. Process - A series of chemical or metallurgical operations conducted to an end. Producer Gases - Obtained by the partial combustion of coal or coke in air. It consists mainly of carbon monoxide and nitrogen, with smaller proportions of hydro- gen, methane, and carbon dioxide. Pulp - A mixture of fine solids and water capable of flowing as a fluid. Its dilution or consistency is specified either as solid-liquid ratio (by weight) or as a percentage of solids (by weight) . Pulp Density - The dilution or consistency of a pulp specified either as solid-to- liquid or as a percentage of solids, both by weight. Pyrite - Iron disulfide, FeS2> often containing small amounts of copper, arsenic, nickel, cobalt, gold, selenium. Pyrometallurgical - Referring to pyrometallurgy involved in winning and refining metals where heat is used, as in roasting and smelting to remove metals from ores. It is the most important and oldest of the extractive processes. Quench - To cool suddenly (as heated steel) by immersion, especially in water or oil. Raf f inate - The solvent-lean, residual feed solution, with one or more con- stituents having been removed by extraction or ion exchange. Reductant - A reducing agent, one which readily parts with valence electrons and by becoming oxidized reduces the acceptor of these electrons. Carbon and hydrogen are important chemical reductants. Reduction - A chemical reaction in which electrons are added to the constitution of the reactant. A reaction which takes place at the cathode in electrolysis. Refractory - A material of a very high melting point with properties that make it suitable for such uses as furnace linings and in kiln construction. Diffi- cult to treat , usually containing a second metallic constituent. Roast - To heat to a point somewhat short of fusing, with or without access to air, so as to expel volatile matter or effect oxidation. In copper metallurgy, applied specifically to heating of the sulfide concentrate in preparation for smelting. Roaster - A device for roasting or a furnace for drying salt cake. A reverbera- tory furnace or a muffle is used in roasting ore. Rotary Drum Filter - The most common type of continuous vacuum filter, having a horizontal drum with a slotted face that turns at 0.1 to 2 rpm in an agitated slurry trough. 13-5 Slaked Lime - A hydrated form of lime, as a dry powder, putty or aqueous sus- pension. The principal constituent of slaked lime is calcium hydroxide. Slurry - Pulp not thick enough to consolidate as a sludge but sufficiently dewatered to flow viscously. Smelting - The chemical reduction of a metal from its ore by a process usually involving fusion, so that the earthy and other impurities, separating as lighter and more fusible slags, can readily be removed from the reduced metal. Thermal processing wherein chemical reactions take place to produce liquid metal from a beneficiated ore. Smelting Furnace - A blast furnace, reverberatory furnace, or electric furnace in which ore is smelted for the separation of a metal. Steam Stripping - Separation of a volatile component from a solution through evaporation produced by the steam's heat. Stoichiometric - A chemical compound, or a batch for synthesis, is said to be stoichiometric when the ratio of its constituents is exactly that demanded by the chemical formula. Stripping - In chemical extraction of minerals, treatment of pregnant solution to remove dissolved values. Sulfonation Roasting - Roasting in which conditions in furnace allow sulfur in feed to recombine with calcined products to form sulfates. Surge Tank - In pumping of ore pulps or solutions, a relatively small tank which maintains at least a minimum inventory of fluid in the system. Tails - The inferior, less valuable, or refuse part of an ore remaining after the valuable constituents have been removed. Tertiary Amine - An organic chemical containing three organic groups attached to nitrogen in its chemical structure. Thickener-Settler - A vessel or apparatus for reducing the portion of water or solution in a pulp. The solids settle to the bottom for removal while the liquid is drawn off the top. Usually large, shallow basins with revolving mechanims to move solids to the discharge. Top Blown Rotary Converter - A furnace, cylindrical in shape, that rotates continuously about its longitudinal axis. The resulting turbulent bath pro- motes good mixing of liquids, solids, and the stream of gas directed, through a lance, at the upper surface of the melt. Triple-Effect Evaporator - Three stages of single-effect evaporation where the vapor from one evaporator is fed into the steam chest of a second evaporator and then into a third and finally to a condenser. The result is an increase of evaporation per pound of steam used. 13-6 Tundish - A rectangular trough lined with fire clay refractories and with one or more refractory nozzles in its base. Underflow Density - The method of describing, usually by weight percent, the ratio of solid to liquid exiting the bottom of a thickener - settler or other concentrating mechanisms. 13-7 14.0 BIBLIOGRAPHY BIBLIOGRAPHY Annotated Bibliography on Processes (1) Agarwal, Jagdish C. and Thomas C. Wilder, 1974, Recovery of Metal Values from Manganese Nodules , U.S. Patent 3,788,831, granted Jan. 29. This invention concerns a method for recovery of copper, nickel, cobalt and molybdenum from deep sea manganese nodules in which the nodules are ground, balled, roasted with reductant gas and leached with an aqueous solution of ammonia and an ammonium salt to obtain a metal rich leachate. (2) Agarwal, J. C., N. Beecher, G.L. Hubred, D.L. Natrig, and R.R. Skarbo, 1976a, Metal Separation bv Fluid Ion Exchange in the Processing of Ocean Nodules , 105th Annual Meeting of AIME, Feb. 22-26. A fluid ion exchange, FIX, process has been developed for recovery of copper and nickel from an ammoniacal carbonate leach liquor of deep ocean nodules. The process is a coextraction of copper and nickel with LIX64N followed by both basic and acid scrubbing of ammonia, selective nickel stripping and copper stripping. Both metals are recovered by electrowinning. The process was selected after a reagent survey, alternative process comparisons, laboratory studies, and continuous operation. Detailed equilibrium data have shown the dramatic effect of free ammonia on the system performance. The process has been piloted and cost estimated. (3) Agarwal, J.C., N. Beecher, D.S. Davis, G.L. Hubred, V.K. Kakaria, and R.N. Kust,^1976b, Processing of Ocean Nodules: A Technical and Economic Review, Journal of Metals, April, pp. 24-31. A detailed comparison of ammoniacal, sulfuric acid and hydrochloric acid processes is presented bv Kennecott researchers, who favor hvdrometallurgical techniques over pyrometallurgical treatments. (4) Anglo American Corporation, 1976, Leaching of Cobalt and/or Nickel Values from Materials , British Patent 1,423,584, granted Feb. 14. According to the invention there is provided a process of leaching cobalt and/or nickel values from a material containing one or both of these metals in an oxidized form including the step of contracting the material with a leach solution containing ammonium sulphite, ammonium bisulphite of a mixture thereof in a substantially non oxidizing atmosphere. 14-1 (5) Anon., 1971, Process Makes Pure Metals from Ocean Nodules, C & EN, May 10, pp. 56-57. Announcement that Deep Sea Ventures has developed a hydrochlorina- tion process for refining sea nodules. A brief description of the process is included. (6) Anon., 1975, Review of the Activities, Edition 18-1975 Manganese Nodules - Metals from the Sea , Metallgesellschaf t AG, Frankfurt am Main, Germany. See separate citations: Kruger et al. 1975 Boin et al. 1975 (7) Anon., 1976, Interocean '76 - 3rd International Conference and Exhibition for Ocean Engineering and Mine Sciences, 15-19 June 1976, Dusseldorf, Germany (in German). See separate citations: Halback et al. 1976 Neuschutz et al. 1976a and 1976b (8) Barbier, J.L., 1973, The Treatment of Nodules, Mines et Metallurgie : Rev. Ind. Min. Metall . , No. 103, p. 25-29 (Trans, from French). Review of possible processes is given on the recovery of Cu, Ni, and Co from deep sea nodules. (9) Bare, C.B. and Joseph ". Pasquali, 1973, Process for the Extraction of Nickel, Copper, and Cobalt from Manganif erous Ores , U.S. Patent 3,751,554, granted Aug. 7,' 1973. This invention describes a process for the extraction of nickel, copper and cobalt from manganif erous ores, such as deep sea nodules, in which the ore is roasted from about 700°F to about 1400°F in a reducing atmosphere. The ore is cooled under non-oxidizing condi- tions, and leached under oxidizing conditions with an ammoniacal- ammonium carbonate solution comprising about 55 to 160 grams per liter of ammonia and about 30 to 120 grams per liter of carbon dioxide. (10) Earner, Herbert E. , Roger N. Kust, and Robert Payne Cox, 1976a, Obtaining Base Metals from Manganese-Containing Ores , German Patent 2,526,388, granted Jan. 2. Improved Cu, Ni, Co, and Mo recovery is achieved from deep sea Mn nodules by leaching the nodules at 35-55° with NH3-containing (NH4)2C03 solution in the presence of > 2 g/1 Cu+ and by continuously regenerating the formed Cu2+ by introduction of CO under 3.5-7 kg/cm^ pressure in the reactor. Leaching is done in series-connected reac- tors. CO is introduced concurrently with the flow of the slurry at 14-2 such a rate as to maintain the desired Cu concentration in each reactor. To obtain constant temperature in the reactors, the heat generated by the reaction is removed through heat exchangers. The MnC03 precipitation is separated from the slurry and the liquid- containing dissolved Ni, Cu, Co, and Mo is extracted to obtain these metals individually. Control of the Cu"^ level in each reac- tor at > 2 g/1 by continuous reduction of Cu'^ improves the effi- ciency of the process. (11) Earner, Herbert E. , David Stanley Davies, Lester J. Szabo, 1976b, Two- Stage Process for Manganese Nodule "eduction or Deoxidation in a Flow- Through Fluidized Bed, German Patent 2,537,940, granted March 11. The reduction of Mn02 in sea-deposit nodules for leaching and recovery of metals is performed by a two-stage process. The dried nodules are calcined and reduced in a fluidized bed at 732° in an oil combustion atmosphere containing 6% 0. Mn oxides in the calcine nodules are further reduced to MnO in a fluidized bed at 732° with a gas containing H20 5.7, CO 42, H 37.8, CO2 2.2 and N 12.3 volume %. The reduced nodules are leached in a fJH3- (NH4)2C03 solution and the recoveries of Cu, Ni, and Co are 90, 90, and 60%, respectively. The metals are separated by ion exchange. (12) Barton, Bruce E. and Paul H. Cardwell, 1974. Fused Salt Electrolysis to Obtain Manganese Metal , H. S. Patent 3,832,295, granted Aug. 27. This invention provides a method for obtaining molten manganese metal by electrolyzing a molten mixture of metal halides comprising a manganese halide and a mixture comprising a halide of a reactant metal, an alkali metal halide and an alkaline earth metal halide. Manganese metal is obtained in a molten phase below the molten mixture of metal halides and elemental halogen is evolved at the anode. The reactant metal is elec trolytically reduced at the cathode. The anode and cathode are both inert and immersed in the molten mixture of metal halides. The halides are preferably chlorides, bromides, or iodides. The reactant metal can be any cathodically reducible metal which can replace manganese from manganese halide in the molten salt bath, but is preferably aluminum or magnesium, (13) Beck, R.R. , and M.E. Messner, 1970, Copper, Nickel, Cobalt and Molybdenum Recovery from Deep Sea Nodules , p. 70-83 in R.P. Ehrlick, Editor, Copper Metallurgy, Metall. Soc. AIME, New York, 371 pp. Deep sea manganese nodules contain interesting amounts of nonferrous metals which could supply future world needs. A smelting process is described for the selective recovery of copper, nickel, cobalt, and molybdenum from manganese nodules. These values are collected and concentrated in a metallic product which is amenable to treatment by subsequent separation processes. 14-3 / (14) Beck, Russell R. , Martin E. Messner, and Charles W. Anderson, 1971b, Smelting of Manganif erous Material , Canadian Patent 871,066, granted May 18. The invention relates generally to the smelting of manganif erous ore material. It is specifically directed to the recovery of non-ferrous metals consisting of copper, nickel, cobalt, and molybdenum from ore materials containing manganese as the primary metal constituent. The invention has particular application in the recovery of copper, nickel, molybdenum and/or cobalt from manganif erous ore materials found on the ocean floor, e.g., from deep sea nodules. (15) Bell, Malcolm Charles Evert, Ferkes Thijs and Ramamritham Sridhar, 1975a, Recovery of Common Metals from Fused Salts , French Patent 2,236,949, granted Feb. 7. A mixture of equal parts of NaCl and KCl, used in refining Ni matte by chlorination and containing metal chlorides, is reclaimed for further use by precipitation the metals through the addition of H„ , H2S, or Al. In an example, the mixture was heated to 800° for 5.5 hr while H2S is bubbled through the melt, at a rate of 0.021 kg/ kg mixture/hour, until 0.115 kg H2S/kg mixture has been added. The metal content of the mixture before and after treatment is: Cu 1.9, 0.19; Ni 1.89, 0.05; Co 0.02, 0.02; Fe 0.7 0.2%. (16) Bell, Malcolm and Ramamritham Sridhar, 1975b, Dressing of Manganese Nodules , German Patent 2,438,773 granted March 20. Mn ores especially nodules containing Mn, Ni, Co, Cu and Fe were treated by reductive roasting at > 1100°. Thus, Mn nodules con- taining Ni 1.14, Cu 0.78, Co 0.19, Fe 5.9, Mn 21.5, Mg 3.16, Ca 1.89, and silicic acid 22.6% were dried at 105°, ground to a sieve size of < 13.3 (DIN 1171), mixed with 12.0% C and 1% S, heated 20 min at 1215° in a gas stream containing H 7.75, CO 7.75, C0„ 15.0 steam 14.3, and N 54.3%, cooled, ground to particle size < 325 mesh, and separated in a magnetic separator. The magnetic fraction, which was 10.5% of the total, contained Ni 12.0, Cu 7.95, Co 2.0, Fe 57.7, and Mn 5.95%. The nonmagnetic fraction contained Ni 0.21, Cu 0.22, Co 0.05, Fe 2.3, and Mn 27.4%. Metal yields were Ni 87.0, Cu 81.0, Co 79.5, Fe 75.0, and Mn 2.5%. (17) Bell, Malcolm C.E. and Ramamritham Sridhar, 1976, Thermal Upgrading of Sea Nodules , Canadian Patent 989179, granted May 18. Generally speaking, the present invention contemplates a process for thermally upgrading a manganif erous oxide ore containing iron and at least one non-ferrous metal value selected from the group consisting of nickel, cobalt, copper and molybdenum. Aggregates of manganif erous oxide minerals containing at least one metal 14-4 value selected from the group consisting of nickel, cobalt, copper iron, and molybdenum, and a reductant in an amount sufficient to insure the reduction of a preponderant part of the metal value and only minor amounts of manganese are heated to a temperature above about 1100°C, e.g., above about 1200°C, to reduce a pre- ponderant part of the metal values to metal and only minor amounts of manganese to metallic manganese and to coalesce the reduced metal values into discrete metallic particles. The reduced aggre- • gates are cooled under conditions to minimize reoxidation of the reduced metal values, and the discrete metallic particles are recovered to provide a concentrate of the metal values. (18) Brooke, J.N. and A.P, Prosser, 1969, Manganese Nodules as a Source of Copper and Nickel - Mineralogical Assessment and Extraction , Transac- tions of the Institution of Mining &. Metallurgy, June 1969, pp. C64- C73, Report of Discussions at Nov. 1969 Meeting, Vol. 79, 1970, pp. C78- C81, and Authors Reply to Discussion, Vol. 79, 1970, pp. C243-C244. A preliminary investigation has been made of the feasibility of the selective extraction of copper and nickel from deep sea manganese nodules. The influence of chemical composition, tex- ture, crystallographic structure and porosity on one method of extraction H2SO4 was studied. The most suitable type of nodule is predicted to be a highly porous AMn02 material with approximately 1 percent of Cu and Ni. The rela- tive suitability of other materials depends on the criteria chosen for assessment; those discussed are recovery, selectivity and pro- ductivity. (19) Brooks, P.T. , K.C. Dian, J.B. Rosenbaum, 1970, Experiments in Processing Marine Nodules , The Proceedings of the IVth International Mineral Pro- cessing Congress, Prague, Czechoslovakia, June 1-6, 1970, pp. 329-333. The lower grade Atlantic nodules were best leached by carbonated ammoniacal solution after reductive roasting. High extraction of manganese, nickel, cobalt, and copper vas achieved with strong leach solution, but considerable difficulty was encountered in preventing premature precipitation of manganese from the pregnant leach liquor. Weak leach solution enabled selective dissolution and recovery of the nickel, cobalt, and copper. The relatively low tenor, calcareous nature, and processing difficulties suggest ' that this type and grade of nodule is not now economically exploita- ble. The relatively higher grade siliceous nodules from the Pacific Ocean were more easily processed by either of two leach methods based on use of recycled sulfur dioxide. High recovery of manganese, nickel, copper, and cobalt as separate marketable products can be achieved. Exploitation might be economically possible provided the nodules could be mined from deep in the ocean and delivered to a plant at low cost. 14-5 (20) Brooks, P.T. and D.A. Martin, 1971, Processing Manganif erous Sea Nodules , Dept. of the Interior, Bureau of Mines Report of Investigation No. 7473, January 1971. Mineral nodules from the ocean floor represent a vast potential source of manganese, nickel, cobalt, and copper. Process devel- opment studies to devise methods for recovering these metals were conducted by the Bureau of Mines using two types of nodules — relatively low-grade calcareous nodules from the Atlantic Ocean and higher grade siliceous nodules from the Pacific Ocean. Ammoniacal carbonate leaching of reduced nodules appeared to have the best potential for processing the calcareous nodules, but the method proved to be impractical for recovering manganese. Extraction of 90 percent or better of the nickel, cobalt, and copper showed the technique to be attractive for recovery of these metals. The siliceous nodules appeared amenda1>le to a process that starts with either a sulfur dioxide leach of a sulfation roast followed by a water leach to dissolve the soluble sulfates. (21) Burns, R.G., 1965, Formation of Cobalt (III) in the Amorphous FeOOH-nH20 Phase of Manganese Nodules , Nature, Vol. 205, March 6, pp. 999. A model and thermodynamics calculations based on the model are made to support the hypothesis that cobalt is in the iron phase of man- ganese nodules. (22) Burns, R.G. and D.W. Fuerstenau, 1966, Electron-Frobe Determination of Inter-Element Relationships in Manganese Nodules . Am. Mineralogist, Vol. 51, May-June, pp. 895-902. The results may be summarized as follows: 1. The iron concentration fluctuates whereas the manganese concentration is relatively uniform throughout a manganese nodule, being slightly higher where iron is low. 2. There is a distinct correlation between Fe, Co, Ti, and Ca. 3. There is a pronounced element coherence between Ni, Cu, Zn, and Mg and between K and Ba. These elements are enriched with Mn and Al in regions of a nodule where the iron concen- tration is low. (23) Caldwell, A. Blake, 1971, Deepsea Ventures Readying its Attack on Pacific Nodules, Mining Engineering, October, pp. 54-55. 14-6 Discussion of Deepsea Ventures plans for mining, processing and marketing nodules. Brief description of hydrochlorination pro- cess is included. (24) Cardwell, Paul. H. , 1973a, Extractive Metallurgy of Ocean Nodules , Presented at the 1973 Mining Convention/Environmental Show of the American Mining Congress, Denver, Colorado, Sept. 9-12. Review of possible processes with emphasis on hydrochlorination process. (25) Cardwell, Paul H. , 1973b, Extractive Metallurgy of Ocean Nodules , Mining Congress Journal, Nov., pp. 38-43 Discussion of the possible processing routes including smelting, sulfation, ammonia leach, and hydrochlorination. (26) Cardwell, Paul H. and James A. Olander, 1974, Method for Separating Nickel from Cobalt in Aqueous Solution , U.S. Patent 3,854,851, granted Dec. 17. A process is provided for separating nickel and cobalt individually from a solution comprising nickel, cobalt and tiianganese halides. The process comprises extracting cobalt halides and some manganese halides with an organic amine at a pH of at least about 3 and separately stripping cobalt and manganese from the extract. Nickel is extracted from the solution with a liquid ion exchange solution. (27) Cardwell, Paul H. , William S. Kane and James A. Olander, 1975a, Nickel - Cobalt Separation from Aqueous Solutions . U. S. Patent 3,894,139, granted July 8. This invention provides a process for obtaining separate solutions of a soluble nickel salt and a soluble cobalt salt from a pregnant aqueous solution comprising a mixture of a cobalt salt and a nickel salt. The nickel and cobalt are extracted together from the preg- nant solution by a liquid ion exchange extraction medium and then selectively stripped from the extract. The pregnant aqueous solu- tion can be derived from the leaching of various ores, such as ocean floor nodule ore, or from, e.g., spent electroplating solu- tions or spent acid pickling liquors, containing nickel, and cobalt. (28) Cardwell, Paul H. , William S. Kane and James A. Olander, 1975b, Method for Separating Nickel from Cobalt , U.S. Patent 3,903,235, granted Sept. 2. The process is provided for separating nickel and cobalt; the pro- cess comprises extracting the nickel and cobalt together with a liquid ion exchange agent. Stripping the nickel from the extractant 14-7 with an acidic aqueous solution and then stripping the cobalt with an aqueous solution at least 6N ■''n hydrogen ion and in chloride ion concentration. (29) Cardwell, Paul H. and William S. Kane, 1976a, Method for Separating Metal Constituents from Ocean Floor Nodules . U.S. Patent 3,950,486, granted April 13. Ocean floor nodules are treated with aqueous hydrogen halide to produce a pregnant leach solution which is subjected to selective solvent extraction to isolate metal values contained in the ore. (30) Cardwell, Paul H. Bruce E. Barton, 1976b, Reduction Method for Pro - ducing Man ganese Metal. U.S. Patent 3,951,647, granted April 20. A process is provided to produce pure elemental metal; such as manganese, by reacting a halide of the elemental metal with aluminum. The halide of the elemental metal is present as a molten pool wherein the temperature at the bottom of the pool is sufficient to maintain the elemental metal molten and the tempera- ture at the top of the pool is below the boiling point of the metal halide in the pool. Aluminum halide is removed as a vapor over- head and a layer of molten metal collects at the bottom of the pool of molten halide. (31) Cardwell, Paul H. , William S, Kane, and Hugh L. McCutcheon, 1976c, Halidation of Ore , British Patent 1,425,845, granted Feb. 18. A process for the refining of ocean floor nodule ore, the ore comprising as primary components the oxides of tetravalent man- ganese and iron and as secondary components, compounds of copper, cobalt and nickel is presented. The process comprises (1) halidating the nodule ore with an aqueous solution of hydrogen halide under reducing conditions to form a pregnant aqueous solution of manganese halide, ferric halide, nickel halide, copper halide and cobalt halide; (2) separating the pregnant aqueous solution from insoluble residue, (3) separating iron value from the remaining halides in solution; and (4) separating the remaining dissolved metal values. (32) Clement, M. , R. Imhof, and E. Mertins, 1973, Contributions to Processing Techniques of Red Sea Type Sediments , Inter-Ocean 73 Conf . , Dusseldorf, Germany, 13-18 Nov., pp. 404-416) (in German). A floatation method to concentrate valuable metals (1% Cu, 5% Zn) from Red Sea sediment is defined. 14-8 (33) Coltrinari, Enzo L. , 1972, Extraction of Nickel from Solutions Containing Nickel and Cobalt Salt s, German Patent 2,152,696, granted May 31. Ni salts were separated from Co salts in ammoniacal solutions by extraction with hydroxy oximes (LIX64) in kerosine and subsequent extraction vjith a H2SO4- (Nll4)2S04 solution of pH 1.3. The Co content in the ammoniacal solution remained constant during extrac- tion. (34) Faugeras, P., 1974, Chemical Treatment of Ocean-Floor Nodules for Industrial Exploitation , Second International Colloquim on Exploita- tion, of the Seas, Bordeaux, France, October 1-4 (Trans, from French). Review of processes, particularly ammonia and sulfur processes. (35) Faugeras, Pierre, Pierre Miquel, and Michele Robaglia, 1975, Obtaining Nickel and Copper from Ores of the Type of Manganese - Containing Nodules_f rom_ Lar^e Marine_Deposits, German Patent 25,07,930, granted Sept. 4 Recovery of Cu and Ni from Mn-containing nodules, such as found on sea bottoms, is accomplished by treating ground nodules with an aqueous H2SO4 solution of pH 2, followed by addition of SO2 at pH 1.5-4 with agitation and subsequent leaching of the mixture with H2SO4 to selectively remove dissolved Cu and Ni. Leaching is doT.ie at pH 1.5-2.0 for 2-12 hr at 35-90°. The solution separated from the solid residue is then contacted with a liquid ion exchanger , such as "LIX" or "KELEX", which will remove Cu from the solution. The Cu content of the ion exchanger is recovered by extraction with a CUSO4-H2SO, solution, followed by electrolysis. The Cli- free solution obtained by the ion exchange is purified and subfje- quently extracted at pH 4 with an organic solvent mixture con- taining di(2-ethylexyl)phosphoric acid 20, BU3PO4 5 and "Solve sso" 75. The 'extraction removes rare earth metals, Zn, and residual Al and Fe, and the aqueous phase will contain the Ni, Co and Mn values. The organic extractant is washed with 1.5N H2SO4 which allows removal of these metals, while Fe remains in the organic phases. The H2S0^ phase is combined with the aqueous phase and Co is removed by extraction with an oxime-containing solvent after pH adjustment to 3.5-4.0. Subsequently, the Co-free solution is used to ob'tain Ni in the form of NiSO, , followed by Mn recovery as Mn02 at [)11 8.5. (36) Fuerstenau, D.W., A. P. Herring, and M. Hoover, 1967, Leaching of Manganese Nodules from the Ocean Floor , paper presented at annual meeting C"f the AIME, Los Angeles, February. Same as paper below entered in AIME Transactions pp. 205-211. 14-9 (37) Fuerstenau, D.W. , A. P. Herring, and M. Hoover, 1973, Characterization and Extraction of Metals from Sea Floor Manganese Nodules , Transactions of Society of Mining Engineers, AIME, Vol. 254, Sept., pp. 205-211. The dissolution of nickel, copper, and cobalt from five different samples of well-characterized deep sea manganese nodules was measured as a function of temperauure, pH, leaching time, and particle size. Nickel, copper, and cobalt extraction each respond to these variables in a different manner. The results of this investigation indicate that copper and nickel can be selectively dissolved from iron and manganese at moderately low pH, provided the reaction is given sufficient time. A high dissolution of cobalt with dilute sulfuric acid can only be achieved from nodules that have a low iron content (5%) . The most probable rate- controlling step for copper and cobalt dissolution is diffusion through the pores of the nodules, and for nickel dissolution chemical reaction appears to be rate-controlling. (38/ Gandon, Louis, Roger Jean and Serge Solar, 1976, Hydrometallurgical Treatment of Polymetallic Ocean Nodules , German Patent 25 30 909, granted Jan. 22. Ni, Co, and Cu are recovered from deep sea Mn nodules by roasting at > 1700° with 3-10% carbonaceous reducing agent, followed by chlorinating at 70-110° in an aqueous medium with CI to obtain water-solution NiCl2» CUCI2 and C0CI2 contaminated with FeCl2. Cu is removed by cementation, followed by removal of Co and Fe by liquid-liquid extraction by using an organic extractant. From the aqueous phase, the Ni is recovered electrolytically , and Co and Fe are removed from the organic phase by extraction with an acid. Thus, 5 kg deep sea nodules containing Mn 26.7, Cu 0.93, Ni 1.15, Co 0.18, and Fe 4.10%, were ground to 0-20 mm and mixed with 5 wt % anthracite. The mixture was then heated to its m.p., and cast in a sand mold. After cooling, the metal phase containing Fe 54.7, Ni 19.6, Cu 13.3, Mn 4.0, Co 2.3, and C 1.88% was granulated to < 500 y. The granules were dispersed in water, and heated to 103°, CI was introduced for 7 hr at 68 g/hr. The resulting solution contained Ni 28.6, Cu 20.3, Fe (total) 85, Co 3.8, Mn 5.4, Mo 1.12 and Fe3+ 0.056 g/1 which corresponded to a 99.7% recovery of Cu, Ni, and Co. (39) Granville, A., 1975, The Recovery of Deep Sea Minerals: Problems and Proi ^pects , Minerals Sci. Eng. , Vol. 7, No. 3, July, pp. 170-188. Some of the properties of f erromanganese nodules and nodule deposits are described in order to provide a basis for an understanding of the developments toward mining the nodules. These developments are outlined, and an attempt has been made to reach a realistic picture of the status and prospects of ocean mining, processing and their economic aspects. 14-10 (40) Halbach, P., K. Koch, H. J. Renner , and W. R.echbug, 1976, Investigations of Pyrometallurgical Processing of Deep Sea Manganese Nodules Using Sewage Sludge as a Reducing Agent , Interocean 76 Conference, Dusseldorf, Germany, 15-19 June, 1976, pp. 87-101 (in German). Sewage sludge is capable of reducing 75-85% of the Cu and Ni present in manganese nodules while manganese remains in the oxide state. (41) Hammond, Allen L. , 1974, Manganese Nodules (I) : Mineral Resources on the Deep Seabed , Science Vol. 183, 8 Feb. 1974, pp. 502-503, Manganese Nodules (II) : Prospects for Deep Sea Mining, 15 Feb. 1974, pp. 644- 646. A short overview of various process routes is included in these two review articles. (42) Han, K.N. , and D.W. Fuerstenau, 1973, Behavior of Metal Ions during Extraction from Ocean Floor Manganese Nodules, First Australasian Conf. on Heat and Mass Transfer , Melbourne, Australia, Sect. 6, pp. 41-48. From the earlier studies of the acid leaching of ccean floor manganese nodules, it was concluded that the minor elements occur essentially mineralogically- independent of the association of the major elements. Two distinctive dissolution behaviors were observed: the dissolution of copper is mainly controlled by pore diffusion while that of nickel is controlled by hetero- geneous reaction and/or pore diffusion depending upon the size of particles. In this investigation, a pore model has been developed for the leaching of copper and nickel which includes both the heterogeneous reactions on the pore surface and pore diffusion. On the basis of this model, the diffusivity of copper was evaluated to be about 2 X 10-10 cm^/sec at pH 1 using nodules ground to a particle size of 100 X 150 mesh. Heterogeneous reaction rate became controlling at pH 1 only when the nodule material had been ground to a diameter of 1 micron or less. (■43) Han, K.N., M. Hoover, and D.W. Fuerstenau, 1974, Ammon ia- Ammon ium Leaching of Deep Sea Manganese Nodules , International Journal of Mineral Processing, Vol. 1, 1974, pp. 215-230. Ammonia- ammonium leaching of samples of nodules from several different locations was carried out after reduction of the nodules under CO/CO2 gas mixtures at 400, 600, and 800 C. In accordance with thermodynamic analysis, nickel, copper and cobalt oxides in 14-11 the nodules are preferentially reduced with a 60/40 gas mixture of CO/CO2. After an initial reduction step with CO/CO2 at 600 C, leaching at room temperature and atmospheric pressure with aqueous ammonia-ammonium carbonate and ammonia-ammonium sulfate solutions yielded high extractions of copper and nickel (80%) , and close to 50% for cobalt. The nature of the pores in nodules from different locations appears to affect the extraction process. A lower reduc- tion temperature is required to obtain the same extraction of nickel, copper and cobalt in a sulfate system than is necessary in a carbonate system. However, a higher manganese content results in the sulfate leaching solutions as compared to the carbonate system, where essentially none of the manganese and iron are extracted. (44) Han, K.N., and D.W. Fuerstenau, 1975a, Preferential Acid Leaching of Nickel, Copper and Cobalt from Ocean-Floor Manganese Nodules , Insti- tution of Mining & Metallurgy, June 1975, pp. C105-109. The preferential acid leaching of nickel, copper and cobalt from sea-floor manganese nodules was investigated as a function of the concentration of sulfuric acid, particle size, temperature and leaching time. The rates of copper and cobalt dissolution appear to be limited by pore diffusion, but nickel by heterogeneous reaction. Consequently, nickel extraction has a strong leaching time-pH and a strong leaching temperature-dependence, whereas copper and cobalt have a strong particle-size dependence. No correlation between cobalt and iron extraction was observed. A selectivity factor was defined to express the extraction of copper, nickel and cobalt over iron and manganese, and conditions by which large values of the selectivity factor can be achieved are discussed. (45) Han, K.N., and D.IJ. Fuerstenau, 1975b, Acid Leaching of Ocean Manganese Nodules at Elevated Temperatures , Int. J. of Mineral Processing 2, pp. 163-171. Sulfuric-acid leaching of deep sea manganese nodules was carried out in an autoclave at controlled oxygen pressure and tempera- ture. The results of this investigation indicate that by raising the temperature to 200 C and the oxygen pressure to 100 psig, it is possible to extract most of the Ni and Cu within a leaching time of one hour without attacking the Fe and Mn phases with markedly reduced acid consumption. These results are discussed in vie ; of an elevated temperature thermodynamic analj'sis. (46) Hanig, G., 1973, High Pressure Sulfuric Acid Leaching of Manganese Nodules , Interocean 73 Conference, Dusseldorf, Germany, 13-18 November 1973, pp. 432-444 (in German). 14-12 This paper deals with the H2S0^ pressure leaching of manganese nodules, one of the many possible processes presently under investigation. Nodules with 1.4% Ni, 1.1% Cu and 0.25% Co originating from the Pacific Ocean were leached in sulfuric acid. It turned out that the grain size and the time of treatment had no significant influence on the leaching behavior. The main factors are the leaching temperature and the amount of surplus acid. At temperatures between 200 and 240 C, 80-90% of the nickel and copper and about 70-80% of the cobalt were recovered, while at a temperature below 100 C, only 60% of the nickel, 40% of the copper, and 10% of the cobalt went into solu- tion. Under the same conditions the solubility of iron is reduced to approximately 2% range due to hydrolysis at the high temperatures. That of manganese is constantly below 5%. (47) Hanig, Gemot, and Micheal J. Meixner, 1974, Drucklaugung von Manganknollen mit Schwelf elsaure , Erzmetall, Vol. 27, No. 7/8, pp. 335-340. Essentially the same information as in Hanig et al. 1973. (48) Herbert, I.C., 1976, Mining Annual Review - Extractive Metallurgy , Mining Engineering Magazine, pp. 263-265. Short review of the progress during 1975 in the mining and pro- cessing of deep sea nodules. (49) Hoover, M. , K.N. Han, and D.W. Fuerstenau, 1975, Segregation Roasting of Nickel, Copper and Cobalt from Deep Sea Manganese Nodules, Int. J. of Mineral Processing, 2, pp. 173-185. The recovery of nickel, copper and cobalt from ocean manganese nodules by a segregation roasting technique was investigated under a wide range of conditions with several nodule samples all of which gave similar results. The best conditions for the segregatioii of the metals were achieved with CaCl2 as the chloride source at a batch retention time of approximately two hours, the highest recoveries were obtained at approximately 850 C for copper and at 1,050 C for nickel and cobalt. At 850 C, copper recovery was 75%, but nickel and cobalt recoveries were only about 25%. At 1,050 C, the nickel and cobalt recoveries were increased to about 60%, but the copper recovery dropped to only 35%. Electron- probe microanalysis showed the segregated metal to be an alloy, indicating that gaseous reactions play an important role in the reduction of chlorides to metal during the process. 14-13 (50) Horn, David R. , 1972, Editor of Papers from a Conference on Ferromanganese Deposits on the Ocean Floor , Lamount-Doherty Geological Observatory, Columbia Univ., Palisades, N.U. , Jan. 20-31, 1972. Several articles of interest. See individual citations. (51) Hubred, G. , 1975, Deep Sea Manganese Nodules: A Review of the Litera- ture , Mineral Sci. Engg. , Vol. 7, No. 1, Jan. 1975, pp. 71-85. A review of the literature on manganese nodules is provided, with emphasis on extractive metallurgy: an overview of other areas such as geochemistry and mining is also provided. (52) lammartino, Nicholas R. , 1974, Metals from Mn Nodules , Chemical Engineering, November 25, 1974, pp. 52-53. Brief review of processing routes is presented. (53) Illis, Alexander and Bernardus J. Brandt, 1975, Selective Process for the Recovery of Metal Values from Sea Nodules , Canadian Patent 974371, granted Sept. 16. A smelting process is described wherein nodules are reduced at 1200''C with CO2/CO mixture added, leached with NH3/CO2 solution after cooling, and Ni , Cu and Co recovered. The residual solids are treated with aqueous SO2 solution to concentrate the Mn in solution and to recover Mn product by further processing. (54) The International Nickel Company of Canada, 1975a, Extraction of Copper , Nickel, Cobalt, or Zinc from an Acid Aqueous Solution Containing One or More of These Metals , Canadian Patent 2,244,004, granted April 11. Extraction of Cu, Ni, Co, or Zn from an aqueous solution con- taining Si02 is done by means of a liquid ion-exchange agent in a water-immiscible organic solvent at pH 4-6 and removing an eventually formed precipitate thereby avoiding the formation of emulsions or foam at the interface. From a Ni-containing solu- tion obtained by leaching ores or wastes with dilute H2SO4, pH 1.8-3.5, Al and Fe are removed by oxidative hydrolysis at pH > 3.0. At pH 4-6, Si02 is precipitated, followed by the extrac- tion of Cu, Co, and Zn with a solution of quaternary Me ammonium thiocyanate in a hydrocarbon solvent. (55) The International Nickel Company of Canada, 1975b, Extraction of Metals from an Aqueous Acid Solution , French Patent 2,264,876, granted Oct. 17. An aqueous acidic solution of 7 g NiSO^/l with impurities of Co 0.5, Cu(II) 0.05, Zn and Fe(III) 0.02 g/1 is extracted for ion exchange by solutions of quaternary ammonium thiocyanates with 14-14 > 1-3 mg/1 SCN .-ind 1.9 moles SCN ions/mole extractable metal is added to the solution which is ti.en extracted with a H2O immis- cible organic solvent. (56) Ito, Hiroshi, Ayaki Nogi, Akitsugu Okuwaki, and Taijiro Okabe, 1975, Hydrometallurgy of Manganese Nodules. II. Leaching of Copper, Nickel and Cobal t from Manganese Nodules with Su lfur ic Acid. Nippon Kagaku Kaishi, pp. 980-984. The leaching tests for nodules heated preliminary at 100 '^-720 C for 3 hr in air were carried out. Results of the leaching of -150 mesh fraction with 0. 5 N H2SO4 showed that Cu in nodules was extracted >80% at 70°, whereas < 10% Co V7as extracted even at 90°. Therefore, Cu, Ni , and Co in nodules could be selec- tively separated by control of leaching temperature. Fe was leached by increasing H2S0^ concentration. Preliminary heating of nodules improved leaching of Co and Mn decreased leaching of Cu, and made the separation among these elements less selective. Leaching with HCl, HNO3 or HCIO4 shovjed almost the same results as those with H2SO4 at 30°. At 90° HCl reduced Mn02 in nodules to MnCl2 (57) Kane, William S. and Paul H. Cardwell, 1973a, Recovery of Metal Values from Ocean Floor Nodules , U. S. Patent 3,752,745, granted Aug. 14. Method for recovering metal values from ocean floor nodules of the type containing iron, manganese, copper, cobalt, and nickel, comprising grinding the nodules into particles, hydrochlorinat- ing to obtain metal chlorides as reaction products, converting the -iron chloride to iron oxide, leaching the reaction products, removing the iron oxide, separating the copper chloride, nickel chloride, and cobalt chloride solutions from the manganese chloride solution by means of liquid ion exchange and recovering the metal values by electrolyzing or by aluminothermal reduction. (58) Kane, William S. and Paul H. Cardwell, 1973b, Processing of Manganese Nodules from the Ocean Floor for Metal Values , U.S. Patent 3,773,635, granted Nov. 20. Method for recovering metal values from ocean floor nodules of the tvpe containing iron, manganese, copper, cobalt, and nickel comprising the nodules into particles, chlorination to obtain vaporized metal chlorides and oxides of carbon or water as reaction products. Condensing the metal chloride reaction pro- ducts, converting the iron chloride to iron oxide while leaching the remaining reaction products with water, recovering the metal chlorides by liquid ion exchange and recovering metal values by electrolyzing. The process is distinguished in its capability of condensing the metal chloride products in fractions. 14-15 (59) Kane, William S. and Paul H. Cardwell, 1974a, Method for Selectively Leaching Metal Values from Ocean Floor Nodules , U. S. Patent 3,795,596, granted Harch 5. Method for recovering metal values from ocean floor nodules of the type containing, iron, manganese, copper, cobalt and nickel comprising grinding the nodules into particles, subjecting the ground nodules to a first leaching step which dissolves copper and nickel, after which the solids are separated, then subjecting the leach liquor to liquid ion exchange and electrolyzing so as to obtain copper and nickel values. Subjecting the solids to a second leach, using ferrous sulfate of ferrous chloride, removing the iron oxide and subjecting the solids to a liquid ion exchange for separating cobalt and man- ganese and then electrolyzing to recover independently the cobalt and manganese values. (60) Kane, William S. and Paul H. Cardwell, 1974b, Mixed Ore Treatment of Ocean T^loor Nodule Ore and Iron Sulfidic Land Based Ores , U.S. Patent 3,809,624, granted May 7. This invention provides a process for simultaneously obtaining the metal values from ocean floor nodule ore and from an iron sulfide mineral. The nodule ore comprises primarily oxides of iron and manganese plus nickel, copper and cobalt compounds. The ore is reacted with the iron sulfide ore, which may also contain desired nonferrous metal values at elevated temperatures in the presence of oxvgen to form iron oxide and the water soluble sulfates of the nonferrous metal. The mixed reacted ores are then leached with V7ater to obtain a solution and the resultant pregnant leach solu- tion is treated, as by liquid ion exchange processes, to obtain separate streams of the individual metal values of nickel, cobalt, copper and manganese. The metals can be obtained by cathodic reduction of the compounds. (61) Kane, William S. and Paul H. Cardwell, 1974c, Method for Separating Metal Values from Ocean Floor Nodule Ore , U. S. Patent 3,810,827, granted May 14. This invention provides a process for obtaining the metal values from ocean floor nodule ore. The ore comprises primarily oxides of iron and manganese plus nickel, copper and cobalt compounds. The ore is treated with an acidic reagent, e.g., SO2, in the absence of oxvgen, to form the water-soluble salt, e.g., the sulfate, of manganese, only. The ore is then leached with water to obtain solution of the manganese salt, which may be further treated to obtain manganese metal. The ore residue can be further treated to extract the other metal values. 14-16 (62) Kane, William S. and Paul H. Cardwell, 1974d, Method of Sulfide Ore and Ocean Floor Nodule Processing , Canadian Patent No. 951, 908 issued July 30, 1974. Sulfide minerals or concentrates containing high concentrates of iron sulfides such as pyrrhotite or pyrite can be mixed with the manganese nodules and roasted in the presence of air or oxygen. The roasting is done at a temperature which will convert the iron sulfide to iron oxide; the sulfides of the other metals contained in the sulfide concentrate such as copper, zinc, and lead are converted to their respective sulfates. The metal oxides and hydroxides contained in the nodules are converted to their sulfates. Thus, upon leaching with water, all of the soluble sulfates are dissolved such as copper and zinc from the sulfide concentrate, and from the nodule manganese, nickel, cobalt, and copper besides other possible minor metal sulfates. Another method of using sulfur dioxide is to react the gas directly with the nodules followed by leaching with water. A third method of using sulfur dioxide is to react the gas directly with dried nodules, followed by leaching with water. (63) Kane, William S. and Paul H. Cardwell, 1974e, Process for Recovering Manganese from It's Ore , U.S. Patent 3,832,165, granted Aug. 27. This invention provides a process for obtaining high purity man- ganese in the molten state from manganese oxide ores containing iroa. The process comprises halidating the ore with a hydrogen halide and leaching to obtain a leach solution comprising man- ganeous and ferric halide and the elemental halogen, extracting the ferric halide from the solution, separating anhydrous man- ganese halide from the solution, reducing the anhydrous manganese halide in the molten state with aluminum to form molten manganese metal and aluminum halide, reacting the aluminum halide with water vapor to form hydrogen halide and aluminum oxides, and recycling the hydrogen halide. Usually where the ore also con- tains a nonferrous metal more noble than manganese, the halide of such nonferrous metal is removed from the leach solution after the ferric halide is removed by precipitation. (64) Kane, William S., Paul H. Cardwell, 1975a, Reduction Mfethod for Separating Metal Values from Ocean Floor Nodule Ore , U.S. Patent 3,869,360, granted March 4. Tnis invention provides a process for obtaining the metal values from ocean floor nodule ore. The ore comprises primarily oxides of iron and manganese plus nickel, copper, and cobalt compounds. 14-17 The ore is treated with an acidic reagent, e.g., sulfur dioxide in the presence of oxygen, to form the water-soluble sulfates of manganese, nickel, copper and cobalt. The reacted ore is leached with water and the resultant leach solution is treated by liquid ion exchange process to form separate solutions of each of nickel, copper, cobalt and manganese salts. These salts can then be cathodically reduced to form the respective metals. (65) Kane, William S. and Hugh L. McCutcnen, and Paul H. Cardwell, 1975b, Recovery of Metal Values from Ocean Floor Nodule Ores by Halidation in Molten Salt Bath , U.S. Patent 3,894,927, granted July 15. This invention provides a process for removing metal values from ocean floor nodule ores comprising contacting the nodule ore with a molten bath of an alkali metal halide and/or an alkaline earth metal halide to form the halides of the manganese, copper, cobalt and nickel present in the ore, and separating the thus formed halides from the reaction mixture, as by vaporization. The mix- ture of halides can then be separated into the individual halides, e.g., by dissolving in water and separating by extraction. Preferably the ore is first dehydrated and the ore can be contacted with a reducing agent to reduce the manganese present to the divalent state. (66) Kane, William S. and Paul H. Cardwell, 1975c, Method of Ocean Floor Nodule Treatment and Electrolytic Recovery of Metals , U.S, Patent 3,901,775, granted Aug. 26. Method of recovering metal values from ocean floor nodules of the type containing iron, manganese, copper, cobalt, and nickel, com- prising grinding the nodules into particles and mixing with a chloride-forming medium so as to obtain metal chlorides as reaction products, including vaporizing the metal chlorides into fractions, converting the iron chloride to iron oxide directly subjecting the manganese chloride fraction to fused salt electro- lysis, leaching an iron oxide-copper chloride fraction, removing iron oxide and electrolyzing so as to obtain copper value, leaching the mixture of nickel, cobalt, and manganese chlorides, and separating the nickel chloride and cobalt chloride by liquid ion exchange and electrolyzing separately, then crystallizing any remaining man- ganese chloride and electrolyzing. (67) Kane, William S. and Paul H. Cardwell, 1975d, Winning of Metal Values from Ore Utilizing Recycled Acid Leaching Agent , U.S. Patent 3,923,615, granted Dec. 2. This invention provides a process for selectively removing desirable metal values from an ocean floor nodule ore wherein the ore is initially leached with an aqueous solution of a mineral acid to 14-18 preferentially form water-soluble salts of nickel and copper which are dissolved in the leach solution, followed by releaching the solid leached ore under reducing conditions using a reducing reagent capable of reducing tetravalent manganese to divalent manganese and of forming water-soluble salts of cobalt and divalent manganese which are in turn dissolved into the leach solution resulting preferably in a single aqueous pregnant leach solution comprising the dissolved salts of copper, cobalt, nickel, and divalent manganese. Preferably, the above four named metal values are individually separated into individual aqueous solutions which preferably are in turn reduced to the elemental metals by electro- lysis procedures. The advantage of this process lies in the fact that it can utilize waste products from this and other industrial procedures. The reducing reagents, which can be used, include ferrous salts obtained from pickling procedures and the mineral acid leach solution can be recycled as spent electrolyte from the electrolysis procedures especially that for the formation of elemental manganese. (68) Kane, William S. and Paul K. Cardwell, 1976a, Two State Selective Leaching of Metal Values from Ocean Floor Nodule Ore , U.S. Patent 3,930,974, granted Jan. 6. This invention provides a process for selectively removing metal values from ocean floor nodule ore by a two-stage procedure. In the first stage, the ore is leached with an aqueous solution of sulfuric acid or a hydrogen halide to preferentially remove nickel and copper. In the second stage, the ore is leached with a reducing agent to preferentially remove cobalt and manganese. The two leach solutions can then be further treated to separate the individual metal values, e.g., by liquid ion exchange pro- cedures. (69) Kane, William S. , Hugh L. McCutchen, and Paul H. Cardwell, 1976b, Direct Recovery of Metals from Fluid Anhydrous Metal Halides Derived from Marine Nodule Halidation , U.S. Patent 3,940,470, granted Feb. 24. This invention provides a process for separating metal values from an anhydrous mixture of metal halides in a fluid state, i.e., vapor or liquid; the separation is made by contacting a fluid mixture containing a halide of at least one less-noble metal selected from the group consisting of iron and manganese, and a halide of at least one more-noble metal selected from the groups consisting of copper, cobalt and nickel, with an ele- mental metal. This procedure is especially effective in the refining of ocean floor nodule ores. 14-19 (70) Kauczor, H.W. , H. Junghanss, and VJ. Roever, 1973, The Hydrometallurgy of Metalliferous Solutions in the Processing of Manganese Nodules , Interocean 73 Conference, Dusseldorf, Germany, 13-18 November, 1973, pp. 469-473 (in German) . A new procedure for the separation and recovery of metals from manganese nodules has been developed which is characterized by the use of ion exchange, precipitation processes and solvent extraction. In the first step copper, nickel and cobalt as well as zinc, iron and aluminum are taken up by the chelate- forming ion exchange resin Lewatit (R) TP 207. Thereby the wanted metals are separated from the excess of manganese. These metals are recovered in the eluate. Its volume is only 12% of the original leach liquor. This makes one of the remarkable items in the whole procedure. Copper which had been fixed on the resin strongly can be recovered directly. During the following steps iron and aluminum are separated by precipitation. Zinc and cobalt are recovered by means of sol- vent extraction methods using tributylphosphate and triisononyl- amine respectively. Nickel remains in the solution and is removed by conventional methods. (71) Kruger, Joachim, and Karl-Heinz Schwarz, 1975, Processing of Manganese Nodules , Review of Activities, Edition 18-1975, Metallgesellschaf t AG, pp. 37-43. Discussion of alternatives in each process step: Leaching (dissolution) , separation and metal recovery. (72) Kruppa, C, 1973, Editor, Proceedings of Interocean '73 , 13-18 Nov., Dusseldorf, Germany, Seehaf en-Ferlag Erik Blumenfeld, Hamburg, 682 p., (in German) . See separate citations: Clement et al . 1973 Supp et al. 1973 Hanig et al. 1973 Ulrich et al. 1973 Meyer-Galow et al. 1973 Kauczor etal. 1973 Caluss et al. 1973 (73) Maekita, Takahiko, Tatsumi Shimzu, and Yoshio Fukuda, 1972, Extraction of Manganese, Nickel, and Other Valuable Metals from Manganese Nodules as Solutions of Water-Soluble Salts , Japanese Patent 72,30,505, granted Nov. 9. 14-20 Powder (100 mesh) Mn nodule 2.5 g was stirred in 10 ml of 5N H2SO4 (diluted with 0-90 ml water) plus 2 ml 30% H2O2 at room temperature for 3 minutes. After filtration, the filtrates contained nearly all of the metals (Mn 100%, Ni > 90%) in the starting nodules. (74) McCutchen, Hugh L. , and Paul II. Cardwell, 1974. Process for the Elec- trolytic Refining of Heavy Metals , U.S. Patent 3,855,089, granted Dec. 17. This invention provides an improved electrolytic process for obtaining the elemental metal by electrodepositing from an aqueous solution of a halide of the metal. The improvement comprises the presence of an alkali metal halide or an alkaline earth meta] halide dissolved in the electrolyte solution. This process is preferably applicable to obtaining the respective metal from the halides of nickel, cobalt, tin, iron, and man- ganese. (75) McCutchen, Hugh, L. , William S. Kane, and Paul H. Cardwell, 1975, Method for Obtaining Metal Values by the Halidation of a Basic Manganif erous Ore with Ferric Chloride Pre-Treatment , U.S. Patent 3,903,236, granted Sept. 2. This invention provides an improvement in the process of obtaining metal values from manganif erous iron-containing ore by halidating the ores and separating the metal values from the ore as the halides. This improvement comprises segregating the iron halides from-the mixed halides and recycling the iron halide to react with the ore prior to halidating. This process is especially applica- ble to ocean floor nodule ores and utilizes preferably hydrogen halides or halogens as the halidating agents, in aqueous solu- tions. (76) Meixner, Micheal , Ulrich Scheffler and Klaus- Herbert Ulrich, 1974, Leaching of Manganese Nodules , German Patent 2,240,586, granted Feb. 28. Mn, Cu, Ni, and Co were extracted from Mn nodules and other areas from the sea bottom by leaching with HCl in gastight vessels at defined HCl-ore ratio, normal pressure, and below the b.p. Mn nodules (1000 g) containing Mn 24, Fe 7.0, Al 3.0, Ni 1.0, Cu 0.9, Co, 024, Mg 2.1, Ca 1.9, Na 2.0, K 1.2, and hydrate water 13% were leached with 1.8 kg HCl (containing 200 g HCl/1) at 0.36:1 HCl-ore ratio for 3 hr at 100° to dissolve Cu 92, Ni 85, Co 20, Mn 20, and Fe 7%. The spent HCl from leaching was recovered after passing through a pyrolysis reaction. 14-21 (77) Menz, Dieter, 1975, Process for Extracting a Manganese Concentrate from Maritime Manganese Or e, U.S. Patent 3,906,075, granted Sept. 16. This invention relates to a process for extracting a manganese oxide concentrate of metallic copper, nickel and cobalt from manganese, such as "maritime manganese concentrations. The ore is first finely ground and then is leached in an aqueous sus- pension at temperatures of 90°-100°C and at a pH value above 3 with sulfur dioxide or sulfurous acid and the resulting man- ganese sulfate is separated by crystallization and subsequent thermal decomposition of the manganese sulfate. The aqueous solution is separated from the residue in a known manner. (78) Mero, John L. , 1965, Process for Separation of Nickel from Cobalt in Ocean Floor Manganif erous Ore Deposits , Patent 3,169,856, granted Feb. 16. This invention relates to a process for recovering nickel, cobalt, and other metals from ocean floor manganif erous ore deposits, and more particularly relates to a process of separating the nickel from, the cobalt found in ocean floor manganif erous deposits. (79) Meyer-Galow, E. , K.H. Schwarz, and U. Boin, 1973, Metal Extraction from Manganese Nodules by Sulfating Treatment , Interocean 73, Dusseldorf, Germany, 13-18 November, pp. 458-468 (in German). Two alternative sulfatizing processes for treating manganese nodules are presented: Sulfatizing roasting Direct aqueous leaching with sulfur-dioxide. The results are based on benchscale test work. The recovery of the valuable metals was determined as a function of some param- eters such as roasting-temperature, grain-size, acid-concentra- tion, retention-time and gas-composition. The further treatment of sulfate leach liquors gained by these processes is schematically summarized. (80) Moreau, Eugene J., Richard S. Opalanko, Norman Schapiro, and Robert E. Lueders, 1976, Manganese Nodules Pelletizing , U.S. Patent 3,942,974, granted March 9. Sodium ions are added to manganese nodules to convert the mont- morillonite clay naturally occurring in the nodule material into a bonding agent for pelletization into spherical agglo- merates. The addition of coking coal to provide bonding strength after reduction, during which the clay bond is destroyed is also disclosed. 14-22 (81) Morgan, Charles L. , 1973, Some Chemical Constraints on the Processing of Manganese Nodules , Marine Research Laboratory, University of Wis- consin-Madison, April. Overview of leaching and metal separation steps with emphasis on the Deepsea Ventures, the Kennecott Copper's and an H2SO4 pro- cesses. (82) Nagaya, Yoshishige, Masaru Fukawa, and Kosuke Murai, 1973, Manganese Nodule Smelting , Japanese Patent 73,56,513, granted Aug. 8. Mn nodules containing Ni, Co, Cu, Mn, and Fe are treated to give a Fe-base alloy of high Ni, Co, and Cu content and a high-C ferro- manganese or silicomanganese by (1) selectively reducing > 90% of the Ni, Co, and Cu oxides in the nodules to metals in a rotary kiln or a shaft furnace, (2) smelting the selectively reduced nodules in an electric or reverberatory furnace to give a Cu-Co- Ni-Fe alloy and a high-Mn oxide slag, and (3) smelting the Mn slag in an electric furnace to recover the Mn values. Thus, green pellets prepared from a 10:3 mixture of ground Mn nodules (Ni 1.09, Co 0.11, Cu 0.68, Fe 8.94, and Mn 28.03%) and anthra- cite powder with addition of 5% pulping waste solution were stirred and heated in a silica tube in the presence of Co at a rate of "^ 10°/min to 900-1100°, then maintained at this tempera- ture for 90 min to effect the selective reduction. The reduced pellets (2.5 kg) were then melted in an alumina crucible at 1500 ± 30° for 60 min to give 0.162 kg metal containing Ni 20.86, Co^2.05, Cu 13.29, Fe 54.01, Mn 5.64, P 2.62, and S 1.50% and 2.59 kg slag containing Ni 0.08, Co 0.009, Cu 0.026, MnO 45.30, and FeO 10.25%. The Mn slag (2.0 kg) was crushed to -3 mm and mixed with coke and quick lime (100:15:15), then melted in a graphite crucible at 1450 ± 30° for 60 min to give 0.796 kg high- C ferromanganese containing Mn 73.80, C 6.67, Si 0.30, P 0.04, S 0.04, Fe 18.93, Ni 0.13, Co 0.02, and Cu 0.06%. The recoveries of Ni, Co, Cu, and Fe in the metallic phase were 95, 93, 97, and 30%, respectively. In the 2nd smelting process, Mn 84, Fe 95, and P 95% were recovered. (83) Neuschutz, D. , U. Scheffler, and H. Junghanss, 1976a, Processing of Manganese Nodules by Sulfuric Acid Pressure Leaching , Interocean 76. Conference, Dusseldorf, Germany, 15-19 June 1976, pp. 102-114 (in German) . Based on sulfuric acid pressure leaching for extraction, a com- plete process flowsheet was developed for the processing of manganese nodules and the recovery of the valuable metals con- tained in them. The major process steps were tested on a laboratory scale. Le .ching tests in which temperature time. 14-23 acid quantity and the type of manganese nodule were varied, have shown that the optimum extractive conditions are 200°C, 3 hours and a residual acid contact of 30 g/1. Under these conditions approximately 95% Ni, 94% Cu, 90% Zn, 67% Co and only about 5% Mn and 1% Fe are dissolved. After partial neutralization with lime, the major quantity of copper contained in the pregnant solution is removed by solvent extraction and recovered by electrowinning. The alternative paths are proposed for further processing of the solution. In one the residual copper and subsequently nickel, cobalt and zinc are separated by chelate-f orming fixed-bed ion exchangers. In the second alternative residual copper remaining after solvent extraction and zinc are precipitated as sulfides with H2S at ambient temperature and processed in a separate cycle. Cost estimates were prepared for the two processes concepts. (84) Neuschutz, D. and U. Scheffler, 1976b, Hydrometallurgical Processing of Red Sea Sediments , Interocean 76 Conference, Dusseldorf, Germany, 15-19 June, 1976, pp. 115-126 (in German). In contrast to most other terrestrial ores, separation of copper and zinc by floatation is not possible at justifiable expense and with a satisfactory metal yield. Hydrometallurgical processing is proposed to extract copper (4%) and zinc (30%) from the con- centrate containing about 80% of minus 5 micron particles and a large amount of water. It is proposed to use sulfuric acid pressure leaching with oxygen and recycled zinc-barren electro- lyte in order to achieve an essentially closed leaching circuit requiring a minimum of chemicals and involving little pollution. After separation of the leach residue, the chloride ions carried over with the washed concentrate will be precipitated with cuprous oxide under reducing conditions in order to achieve the values permissible in the electrolytic production of zinc. Copper is recovered by solvent extraction and electrowinning. The zinc- bearing raffinate is concentrated by evaporation, purified and fed into the electrolysis circuit. Based on test carried out, a cost estimate of a process is presented. (85) Okabe, Taijiro and Akitsugu, Okuwaki, 1974a, Treatment of Manganese Nodules , Japanese Patent 74,123,110, granted Nov. 25. Mn nodules from the sea bed are treated with NH4 salts or reducing sulfuric acids to extract Cu, Ni, and Co as the amine complexes. Nodules (5 g) < 105 u particle size, containing ^In 23.27, Cu 0.76, Ni 0.87, and Co 0.28% were treated with 150 ml 5M NH3-2M (NH4)2S03 in an autoclave at 60° for 2 or 8 hr; extraction was (1,83,82,85), respectively. 14-24 (86) Okabe, Taij iro and Akitsugu Okuwaki, 1974b, Treatment of Manganese Nodules , Japanese Patent 74,123,111, granted November 25. Nodules are treated with dilute H2SO4 containing an organic compound. Mn nodules (25 g) < 105y particle size containing Mn 23.27, Cu 0.76, Ni 0.87, and Co 0.28% was stirred with 400 ml 2.5M H2SO4-I.5M MeOH at 60° for 2 hr; extraction was Ni 94, Cu 96, Co and Mn 100% each. HCHO , dextrin, pulp, and activated sludge were also used. (87) Okajima, Yasuhio, 1975, Treatment of Manganese Nodules , German Patent 2,522,969, granted Nov. 27. Mn nodules, dried at 105° and containing Mn 26.0, Fe 10.3, Ni 1.12, Co 0.14, Cu 0.62, Si02 13.8, AI2O3 3.7, MgO 2.7, and CI 0.24 wt %, were ground with 7 wt % coal to particles of < 0.59 mm size. The mixture was then pelletized to 5-20 mm diameter, the pellets were roasted in a shaft furnace with combustion off- gases containing 0-0.7 vol. % O2 at 550-750° for 30-60 min. The reduced pellets were cooled to 100-400°, then wet ground to particles of < 0.21 mm size, followed by addition of 0.25 wt % NaCl as an aqueous solution of 50 g/1 concentration. The mix- ture was then slurried with aqueous (NH4)C03 in a solids liquid ratio of 1:6 to obtain a leach liquor containing NH3 50 and CO2 30 g/1. The slurry was contacted with air for 30 min at the rate of 3 l./min for 3 hr. After separation of the solid residue, the liquid phase containing Ni 7.8, Cu 4.3, Co 0.90, Fe 0.035, and Mn 1.6 g/1 corresponding to Ni 96, Cu 97, and Co 90% recovery, was treated with an aqueous Na2S03 solution containing 50 g Na2S03/l. and with air at the rate of 2 l./min to pre- cipitate residual Fe and Mn. After this treatment both the Fe:Ni and the Mn:Ni ratios were < 0.0001. The filtrate was then treated with an aqueous Na2S solution of 100 g/1. concentra- tion for 30 min to precipitation Co and Cu. The filtrate of this treatment was used for Ni recovery. The solid leach residue, after combining it with the precipitate from the Na2S03 treatment was used for making f erromanganese containing Mn 43 and Fe 16%. The f erromanganese can be readily enriched for metallurgical applications. (88) Pasho, D.W. and J. A. Mcintosh, 1976, Recoverable Nickel and Copper from Manganese Nodules in the Northeast Equatorial Pacific-Preliminary Results , CIM Bulletin, September 1976, pp. 15-16. The resource estimate of manganese nodules has been exaggerated in the literature. Referring to an area in Northwest Equatorial Pacific only and defining a mining site as capable of yielding an annual production of 3 million dry metric tons of nodules for 20 years. There is a 50 percent chance that there are more than 30 mine sites that may become commercially exploitable before the year 2000. 14-25 (89) Pearson, John S,, 1975, Ocean Floor Mining , Noyes Data Corporation, Park Ridge, New Jersey, 202 p. Contains a review of mineral extraction processes pp. 80-106. (90) Pemsler, J. Paul and John K. Litchfield, 1976, Solvent-In-Pulp Extrac - tion of Copper and Nickel from Ammoniacal Leach Slurries , U.S. Patent 3,950,487, granted April 13. A process in which copper, nickel, cobalt and molybdenum are recovered by leaching comminuted raw manganese nodules with an aqueous ammoniacal leach solution containing cuprous ions. An improvement is disclosed in which the metal values are ex- tracted directly from the leach slurry with an organic extractant. To accomplish such extraction the amount by weight of solids in the slurry is maintained at less than 20 percent, the pH of the slurry is lowered to 9.5, the volumes of slurry and organic are maintained at a ratio so that the organic is the continuous phase and the organic and slurry are mixed with gentle agitation. (91) Pemsler, Joseph P., and Litchfield, John K. , 1976a, Obtaining Base Metals from Manganese-Containing Ores , German Patent 2,526,395, granted Jan. 1976. Base metals are recovered from Mn nodules without excessive loss of organic extractants or the use of ion exchangers by mixing the ore with a solution of Cu(I) vjhich reduces Mn oxides and solubilizes valuable metals with formation of Cu(II) , reducing Cu(II) , passing a reducing gas into the reactor, and extracting the resulting slurry at pH < 10.0 with sufficient organic extractant to form a continuous organic phase. Thus, a Cu(I)- treated, 17% solid slurry with Cu, Ni, Co, and Mn content of the liquor 3.50, 4.48, 0.036, and 9.85 g/1. and of the solids 0.316, 0.466, 0.0722, and 0%, respectively is treated at 40° with 45 g/1. NH3 and 44 g/1. CO2 and extracted twice with 1 vol. LIX64N solution in Napoleum, resulting in Cu, Ni, and content in the first organic extract 3.52, 2.99 and 0.010 g/1., and the second organic extract 0.12, 1.16 and 0.013 g/1. The total extraction Cu, Ni, and Co is 94, 82, and 18%, respectively and the loss extractant is 140 ml/1000 1. slurry. When the slurry is treated with NH3 84 and CO2 38 g/1. resulting in pH 10.7, in extractant loss is 950 ml./lOOO 1. (92) Queneau, Paul Etienne, Herm Jan Roorda, and Stanley Charles Townsend, 1975, Chlorine Leaching of Non-Ferrous Metal Values with Ammoniacal Solutions, U.S. Patent 3,880,651, granted April 29. 14-26 U A ferruginous material containing at least one non-ferrous value selected from the group consisting of nickel, cobalt, copper and manganese is, after reduction, immersed in water and then con- tacted with gaseous chlorine until the pH value of the immersion decreases to below 4 and sufficient amounts of ammonia or ammonia- containing compounds are added with the chlorine to maintain the pH value between about 4 and 2 so as to dissolve a preponderant part of the non-ferrous value is recovered from the resulting solution. The ferruginous material can be an oxide ore including sea nodules, or a roasted sulfide ore that has been reduced, ad- vantageously selectively reduced. (93) Redman, M.J., 1973, Extraction of Metal Values from Complex Ores , U.S. Patent 3,734,715, granted May 22. Copper, nickel, cobalt and molybdenum may be extracted from com- plex ores containing manganese, iron, copper, nickel, cobalt and molybdenum by subjecting the complex ore to gaseous reductants at temperatures in the range of 300 to 800 C and leaching the reduced complex ore with a solution of an ammonium salt in aqueous ammonia. (94) Roever, Wilhelm, Helmut Junghanss, Alfred Roeder, Hans-Werner Kauczor, and Gunter Kuhne, 1975, Process for the Separation of Copper from the Metals Iron, Cobalt, Nickel and/or Manganese , Canadian Patent 965,960, granted April 15. This invention describes a process for the quantitative separation of c/3pper from the bivalent metal cobalt, nickel, iron and/or manganese from aqueous solutions by means of ion exchange, charac- terized in that a mineral acidic solution at a pH value of 1.3- 2.5 is passed at a temperature, in the range of 25-80°C, over at least two series-connected exchanger columns containing a com- plex forming ion exchanger resin with aminocarboxylic acid or iminodicarboxylic acid groups as active groups and that the first series-connected exchanger column, when only loaded with copper, is disconnected and eluted as soon as copper breaks through in the last column. (95) Rolf, Ramon F. , 1969, Process for the Selective Recovery of Manganese and Iron from Ores , U.S. Patent 3,471,285, granted Oct. 7. This invention provides a process for the selective separation of manganese and iron values from ores, such as the manganese sea nodules, which likewise contain significant quantities of nickel and cobalt values. The separation is accomplished by first reducing the ore at high temperatures and leaching the reduced ore with ammonium sulfate to concentrate the manganese and iron values in the liquid phase. 14-27 (96) Sandberg, Richard G. and Paul H. Cardwell , 1975, Separation of Non - Ferrous Metals from Manganese Oxide Ores , German Patent 2,522,954, granted Dec. 11. Non-ferrous metals, such as Mn, Co, Ni, and Cu , are recovered from oxide Mn ores, for example deep-sea Mn nodules, by grinding the ore to 1.7 mm size and reducing with a carbonaceous reductant followed by treatment with FeCl3, AICI3, TiCl4, or SiCl4 at 50- 160° to convert the non-ferrous metals to leachable halides. Reduction and halogenation can proceed simultaneously and leaching of the halogenated ore is done at pH < 3. Thus, deep-sea Mn nodules containing Mn 29 and Fe 6 wt % and smaller amounts of Cu, Co, and Ni were ground to < 0.5 mm and dried for 2 hr at 330°. Then 200 g of the dried ore were mixed with 50 g coal of < 0.15 mm and the mixture was placed in a furnace and flushed with nitrogen at 150 ml/min, and heated at 500° for 2 hr. The reduced ore was charged to a dry porcelain ball mill and heated with anhydrous AICI3 to 140° for 90 min. The cooled product was leached for 15 min with an aqueous acid solution of pH 2, and filtered. The filtrate contained 95% of the Mn content of the nodules and 100% of the Cu. The Ni and Co recoveries were 97%. (97) Scheffler, Ulrich and Klaus-Herbert Ulrich, 1974, Dressing of Manganese Nodules , German Patent 2,249,302, granted April 11. In the dressing of Mn nodules, the chlorination yield was improved by a HCl-Cl-steam-0-C02-N mixture, whereby the chlorine bonded to worthless components was recycled. Thus, Mn nodule particles (< 3 mm) were treated at 600° in a fluidized bed by a gas con- taining HCl 5.4, CI 2.5, steam 39.0, 0.6, and CO2 + N 52.2 vol.% at 5% CI excess with respect to the stoichiometric amount for the chlorination of the Cu 98, Ni 95, Co 95, Mn 60, and Mg 80%. The slurry obtained after separation of chlorinated Cu, Ni, Co, and Mn containing 70 kg chloride (MgCl2) in 100 kg water was treated bv pyrolysis for CI recovery. (98) Schobert, Harold H. , Roger C. Field, and Paul H. Cardwell, 1976, Reduc - tion to Manganese Metal Using Metal Transporting Compounds , U. S. Patent 3,950,162, granted April 13. This invention provides a process for obtaining substantially pure manganese metal by the reduction of a manganese oxide of a man- ganese halide by reaction with a subhalide of a transport metal. The transport metal can be aluminum, silicon, or titanium. In preferred embodiments of this procedure, a continuous closed cycle process is carried out wherein the transport metal is reconverted to its subhalide and recycled for reaction with additional manganese compound. 14-28 (99) Scharwz, Karl-Heinz, and Udo Boin, 1974, Verarbeitung von Mangan - knollen unter Einsatz schwef elhaltiger Reagenzien , Erzmetall, Vol, 27, No. 7/8, pp. 341-345. See Meyer-Galow, 1973, for essentially same results. I (100) Sisselman, Robert, 1975, Ocean Miners Take Soundings on Legal Problems , Development Alternatives , Engineering and Mining Journal, April, pp. 75-86. Some description information on each of the major processes. (101) Skarbo, R.R., 1973a, Extraction of Copper and Nickel from Manganese Nodules , U.S. Patent 3,723,095, granted March 27. This invention relates to a process for selectively recovering copper and nickel from complex ores containing manganese, iron, copper, nickel, cobalt, and molybdenum. It has been discovered that copper and nickel are selectively removed from complex ores containing iron, manganese, copper and nickel by leaching the ore with an aqueous solution con- taining the manganous ion (Mn++) . Substantially all of the copper and nickel are thereby solubilized and pass into the aqueous leach phase. The copper and nickel may then be recovered from the aqueous leach liquor by an desirable method. (102) Skarbo, R.R., 1973b, Extraction of Metal Values from Manganese Deep Sea Nodules , U.S. Patent 3,728,195, granted April 17. Copper, nickel, cobalt and molybdenum are extracted from deep sea manganese nodules by the process of this invention which comprises the steps of charging manganese deep sea nodules and a leach solution of aqueous ammonia and an ammonium salt or sodium chloride to a pressure vessel or an autoclave to provide a liquid- solids suspension and heating the liquid-solids suspension to effect reduction of the tetravalent manganese phase of the nodules whereby the nickel, copper, cobalt and molybdenum in the deep sea manganese nodule report to the leach solution. The metal values 14-29 may then be recovered from the leach solution using conventional technology. In another embodiment of the invention a reducing gas is introduced prior to heating the liquid-solid suspension. In this process the reduction of the tetravalent manganese phase is predominantly affected by the reducing gas used. (103) Skarbo, Ronald R. , 1974a, Separation of Nickel from Salt Mixtures Containing Nickel and Other Metal Salts , U.S. Patent 3,798,307, granted March 19. A new method for the separation of nickel from a metal salt mix- ture is disclosed. The method involves dissolution of the salt mixture at an elevated temperature in a solution of ammonium hydroxide and ammonium chloride, rapidly cooling the solution under controlled condition of pH and ammonium chloride concen- tration to selectively separate nickel as crystalline nickel ammine chloride. (104) Skarbo, Ronald R. and Antonio H. Miguel, 1974b, Metal Recovery from Salt Mixtures of Copper and Metals More Electropositive than Copper , U.S. Patent 3,810,971, granted May 14. Copper may be efficiently separated from metal salt mixtures, i.e. carbonate, chloride and sulfates by a process that comprises dissolving a portion of the metal salt mixture in hydrochloric acid, reducing the other portion of the metal salt mixture with a reduced portion along v\7ith elemental sulfur to precipitate out substantially all of the copper in the metal salt mixture as cuprous sulfide. (105) Skarbo, Ronald R. , 1974c, Cobalt Stripping from Oxime Solutions , U.S. Patent 3,849,534, granted Nov. 19. Cobalt is removed from an oxime solution by contact with hydro- chloric acid-sodium chloride solution for 30 minutes at 40-60°C. During contact the cobalt is released by the oxime to the acid solution where it remains as the organic and acid phases are allowed to separate. (106) Skarbo, Ronald R. , 1974d, Selective Stripping Process , U.S. Patent 3,853,725, granted Dec. 10. The present invention relates to an acid stripping process for recovering nickel selectively from an organic oxime containing copper and nickel. By employing only limited am.ounts of acid, sufficient to provide about equivalent latent hydrogen ion therein to latent nickel ion of the oxime stream, nickel can be stripped in a multistage counter-current process preferentially to copper. Hydrogen ion of the acid exchanges with nickel ion 14-30 u of the oxime molecule in accordance with the equation: (R-Ni) + 2(h"'") ^ (2R H) + (Ni'^)aqu 2 org aqu org Wherein R is the oxime organic moiety and R2Ni is a chelate of nickel and the oxime. The acid selectively strips nickel when the mole ratio of exchangeable hydrogen ion of the incoming fresh acid to nickel ion of the incoming fresh oxime is maintained between about 1.8 and 2.2. (107) Skarbo , Ronald R. , 1974e, Process for Recovering Nickel Selectively , U.S. Patent 3,855,090, granted Dec. 17, German Patent 2,333,003. In a process wherein copper and nickel are recovered separately from an oxime extractant by separate stripping and electrowinning steps, nickel concentration in the copper electrowinning step is maintained below about 10 g/1 by directing a portion of a metal- barren acid stream from the copper electrowinning step to the nickel stripping step. Low copper concentration in the aqueous stream from the nickel electrowinning step is effected by main- taining the mole ratio of exchangeable hydrogen ion of the acid to nickel ion of the oxime between 1.8 and 2.2 in the nickel stripping step in accordance with the equation: (R°„Ni) +2(h''") ^ (2R°H) + (Ni''"'') 2 org aqu org aqu wVferein R is the oxime organic moiety and R2Ni is a chelate of nickel and the oxime. (108) Skarbo, Ronald R. , Robert Edward Lueders, Edmond Arthur Morin, and Hal Douglas Peterson, 1974f, Separation of Ammonia from Organic Solutions of an Oxime-Metal Chelate , German Patent 2,360,147, granted June 27. In the selective separation of NH3 from the orr;anic solution to obtain an organic effluent containing < 50 mg/1. NH3 , the organic solution is contacted counter currently with an aqueous solution of (NH4)2C03, NH^HCO^, alkali hydrogen carbonate, or mixtures thereof, so that no emulsion is formed between the aqueous and the organic solutions, the aqueous solution having a lower NH3 concentration than the equilibrium concentration. The effluent organic solution with the reduced NH3 content is then separated from the aqueous extracting solution and the residual NH3 content is preferably separated by contacting the aqueous solution with an aqueous H2SO/ solution having such a pH that no metal is ex- tracted from the organic solution. Alternatively, the residual NH3 can be removed from the organic solution by extracting with H2O. 14-31 (109) Skarbo, Ronald R. , Robert E. Lueders, Edmond A. Morin, and Hal Douglas Peterson, 1974g, Elimination of Ammonia from Organic Solutions of a Metal-Oxime Chelate , French Patent 2,211,453, granted July 19. The selective removal of NH3 to < 50 mg/1. from an organic solu- tion containing NH3 and a metal oxime chelate is achieved by a counter-current treatment with an alkali metal or ammonium car- bonate or bicarbonate in aqueous solution. The NH3 concentration in the solution is maintained at a level below that which would be in equilibrium with 50 mg/1. in the organic phase, and no noticeable emulsif icat ion occurs between the aqueous and organic phases. Practically all the remaining NH3 can be removed by sub- sequent treatment with H2SO4. The method has been applied to solutions containing 2-hydroxy-4-nonylbenzophenone oxime chelate with Ni and/or Cu. In one example, a continuous countercurrent circuit consisting of mixer and separator was used to separate NH3 1.01 g/1. by using M NH4HCO3. The contact time was 2 min and the temperature was 41°. After 3 hr and 10 min the NH3 content of the organic phase was 0.038 g/1. and no appreciable quantity of Cu or Ni had been leached into aqueous phase. (110) Skarbo, Ronald, Water E. Galin, and David L. Natwig, 1975a, Cobalt Stripping from Oximes, U.S. Patent 3,867,506, granted Feb. 18. Cobalt is removed from an oxime solution by contact with a mixture of sulfuric acid and glacial acetic acid or lower alkanol. During contact the cobalt is released by the oxime to the acid solution where it remains as the organic acid phases are allowed to separate. (111) Skarbo, Ronald R. , 1975b, Nickel Extraction and Stripping Using Oximes and Ammoniacal Carbonate Solutions , U.S. Patent 3,907,966, granted Sept. 23. Ammoniacal solutions containing nickel and other metal values are adjusted to low NH^ concentration and treated by liquid-liquid ion exchange extraction with an oxime organic extractant and then the loaded organic extractant is stripped with a highly concen- trated ammonium salt solution to yield a strip concentrate from which nickel or nickel oxide is recoverable without electrolytic treatment. (112) Skarbo, Ronald R. and David L. Natwig, 1976a, Process for Recovering Copper from Basic Carbonate Mixtures , U.S. Patent 3,950,488, granted April 13. Method for treating basic salts to preferentially extract copper without using a base to maintain the required pH. The liberation of unreacted hydrogen ions during the ion exchange reaction is eliminated by partially dissolving the basic salt mixture with an 14-32 amount of acid that is equivalent to the non-copper content of the basic mixture. An amount of organic ion exchange extractant is selected so that the exchangeable hydrogen ion on the extractant is equivalent to or in excess of the amount of copper in the metal bearing liquor. (113) Skarbo, Ronald R. , 197fb, Selective Solvent Extraction Process for Copper , British Patent, 1,445,123, granted Aug. 4. This invention relates to a solvent extraction process for recovering copper selectively from high pH solutions con- taining nickel and copper. The present invention is based upon the discovery that when an ammoniacal leach solution having a pH above 9 and containing copper and nickel is admixed with an organic stream comprising an oxime in a water-immiscible organic solvent, nickel concentration in the organic stream, as a function of copper concentration in the aqueous stream, reaches a maximum and then reduced to about zero as the copper concentration in the organic stream is increased to approach the capacity of the oxime for retaining metal ion. Furthermore, in the same system, the concentration of copper in the organic stream in equilibrium with copper in the aqueous stream does not pass through a maximum and then decreases as a function of copper concentration in the aqueous stream but continues to increase up to the capacity of the oxime for retaining metal ion. This phenomenon is observed regardless of whether the leach solution contains other metals suc*h as cobalt. Thus controlling the concentration of copper in the organic stream, it is possible to exclude nickel from the organic stream. (114) Sridhar, R. , 1973 & 1974, Thermal Upgrading of Sea Nodules , Talk delivered at AIME, 1973, Dallas, Texas and reported in J. Metals, Dec. 1974, pp.l8- 22. Description of INCO's pyrometallurgical process for sea nodules. Some bench scale experimental results reported, (115) Sridhar, Ramamritham, John Stuart, and Malcolm Charles Evert Bell, 1975, Recovery of One or More Non-Fcrrous Metals from Manganese - Containing Oxides , German Patent 2,510,243, granted Sept. 25. Nonferrous metals, such as Ni, Cu and Co, are recovered in high yield from Mn-containing ores, for example deep sea nodules. The nodules are reduced at 800-1000° with a carbonaceous reduc- tant, then melted at 1250-1600° to obtain a molten metal phase containing Ni , Cu, Co, Fe, and a slag of high Mn and Si content. After separation of the slag, the metal phase is blo\>m with an oxygen-containing gas to reduce the Mn and Fe contents by slag formation and the metal phase is then sulfidized by addition of 14-33 elemental S or pyrite. This treatment converts the Cu, Co and Ni content to a matte of CU2S, NiS, and CoS from which these metals can be recovered by conventional methods, for example by carbonyl formation. Thus, deep sea nodules containing Ni 1.14, Cu 0.78, Co 0.19,, Fe 5.9, Mn 21.75, Si02 22.6, Mg 3.16, and Ca 1.89% dry wt were reduced at 1000° with 8.12% coke. The selec- tively reduced nodules were then melted at 1425° and a metal phase constituting '1^8% of the total nodule wt and containing Ni 14.5, Cu 9.3, Co 2.3, Fe 70.2, Mn 2.5, and C 0.74% was obtained. The slag generated contained Ni 0.045, Cu 0.058, Co 0.009, Fe 0.88, Mn 27.2 and Si 37.5%. The metal phase was then blown with oxygen-containing gas and 3% sulfur was added and blowinp with the oxidizing gas was continued until the Fe content was reduced to < 5%. The substantially Fe-free, sulfidized metal was then cooled, granulated, and its nonferrous content recovered by conventional methods. (116) Sridhar, R. , Jones, W. E., and Warner, J, S., 1976, "Extraction of Copper, Nickel and Cobalt from Sea Nodules," Journal of Metals, April 1976, pp. 32-37. Description of INCO ' s pyrometallurgical process for sea nodules. Some bench scale experimental results reported. (117) Supp, Ach, R. Nebe, and H. Schoner, 1973, Physical Processing of Red Sea Type Sediments , Interocean 73 Conference, Dusseldorf, Germany, 13-18 November, 1974, pp. 417-431 (in German). The authors present a flow-sheet for concentrating the Red Sea ores derived from the floatation tests that were carried out, (118) Swan, D. A., 1974, The Potential of Manganese Nodules as a Future Mineral Resource, Marine Technology, Jan., pp. 9-18. No process information. (119) Szabo, Lester J., 1975b, Obtaining Metals from a Manganese-Containing Ore , German Patent 2,524,598, granted Dec. 11. Recovery of Ni, Co, and Mo from deep sea Mn nodules by the Cuprion process is improved by maintaining in the leaching system a Cu concentration of > 10 g/1. The increased Cu concentration can be achieved by introducing a reducing gas, such as a mixture of CO and H, in the leaching system and by recycling a portion of the reduced solution to the leaching system after separation of precipitated basic carbonates. Reducing conditions in the leaching system allow the use of smaller leaching vessels or conversely the use of fewer large-capacity vessels. 14-34 (120) Szabo, Lester John, 1975a, Recovery of Metal Values from Manganese Deep Sea Nodules , South African Patent 8885/73, granted on Feb. 7. This invention relates to the recovery of metal values from man- ganese deep sea nodules, and more particularly to the recovery of copper, nickel, cobalt and molybdenum by the direct leaching of comminuted raw manganese nodules with an aqueous ammoniacal leach solution containing cuprous ions. (121) Szabo, Lester, J., 1976, Recovery of Metal Values from Manganese Deep Sea Nodules using Ammoniacal Cuprous Leach Solution , U.S. Patent 3,983,017, granted Sept. 28. A process in which copper, nickel, cobalt and molybdenum are recovered by direct leaching of comminuted raw manganese nodules with an aqueous ammoniacal leach solution containing cuprous ions. During the process cuprous ions reduced the nodules, are consumed, and are oxidized to cupric ions. The cuprous ions are regenerated by reducing the cupric ions with a reducing gas. (122) Taylor, Donald M. , 1971, Worthless Nodules Become Valuable , Ocean Industry, June, pp. 27-28. Announcement that Deepsea Ventures has developed a hydrochlorina- tion process for sea nodules. A brief description of the process is included. (123) Tinsley, C. Richard, 1975, Processing - No Longer a Problem . Mining Engineering, April 1975, pp. 53-55. Review of possible processing routes. (124) U.S. Dept. of the Interior, Bureau of Mines, 1967, Summary of the Bureau of Mines , research to extract manganese and other metals from undersea nodules. Salt Lake City. See Brooks et al. (1971) - Bureau of Mines Report of Investigation No. 7473. (125) Ulrich, K.H., U. Scheffler, and M.J. Meixner , " 1973 , Processing of Man- ganese Nodules by Acid Leaching , Interocean 73 Conference, Dusseldorf, Germany, 13-18 November, pp. 445-457 (in German). Investigations of the processing of manganese nodules from the Pacific Ocean involve leaching tests with either hydrochloric or sulfuric acid. Leaching temperature was varied between 25°C and 250°C. Between 200°C and 250''C copper and nickel were ex- 14-35 tracted with an efficiency of more than 85% up to 98%. Less than 2% of the iron content was dissolved, which is very important to minimize the costs of the following pregnant solution-treatment. Manganese extraction decreasing during leaching tests under oxidizing conditions with increasing leaching temperatures from about 40% to about 5%. During sulfuric acid leaching in any case manganese dissolution from manganese dioxide is very small because of the oxidizing conditions. The degree of cobalt extractions depends strongly on the specific acid addition and varies from about 30% to 60%. Flow sheets for leaching operations and regeneration of barren solution are presented. Operating costs are estimated. Pressure leaching with regeneration of hydrochloric acid is recommended for processing plants were problems will exist with the deposition of residues and the drain off of barren solu- tions. Without these problems an economic way to process man- ganese nodules is pressure leaching with sulfuric acid. (126) Van Heck, M.S. and R.W. Barlett, 1973. Kinetics of Sulfation of Atlantic Ocean Manganese Nodules . Metallurgical Transactions, 4, April, pp. 941-947. The sulfation kinetics of manganese nodules from the Blake Plateau of the Atlantic Ocean were determined using a TGA method in low concentrations of SO2 typical of power plant stack gases. Above 400°C the rate is proportional to the SO2 pressure and the unreacted solid fraction. The rate is inde- pendent of nodule particle size and the apparent activation energy is low. The sorption capacity is complete sulfation of the major constituent oxides; Mn , Fe, and Ca. Leaching com- pletely sulfated nodules in boiling water provides Ni, Cu, and Co extractions above 80%. Much of the manganese but little of the iron is also extracted. At 300°C the SO2 sorption capacity is lower with most of the sulfation attributed to the manganese oxides. (127) Vasil'chikov, N.V., G.B. Shirer, Yu. A. Matsepon, I.F. Krasnykh, a E.A. Grishankova, Iron-Manganese Nodules from the Ocean Floor - Raw Materials for the Production of Cobalt, Nickel, Manganese and Copper , Tsvet. Metal. 1968, 41(1;, 40-2 (Russian translation to English). A sample of nodules from the Pacific Ocean bottom was examined by mineralogical , chemical, spectrographic , and X-ray methods. The nodules contained Mn 24.2, Fe 6.5, Ni 1.2, Co 0.25, Cu 1.4, V2O5 0.03, and M0O3 0.07%. The Ni, Co, and Cu were apparently absorbed by the Mn-Fe hydroxides from sea water. Reducing smelting of the nodules gave recovery in the slag of Mn 84.5 and Pe 2.8% and in the metal of Mn 8.7, Fe 97, Ni 100, Co 93.5, and Cu 86%. 14-36 (128) Wakefield, B.D. , 1971, Deepsea Nodules Yield Minerals , Iron Age, May 13, pp. 74-76. (129) Weston, David, 1974a, Hydrometallurgical Treatment of Nickel, Cobalt , and Copper Con taining Materials, U.S. Patent 3,793,430, granted Feb. 19. A process for the recovery by leaching of nickel, cobalt, and copper using sulfuric acid and employing agents capable of introducing alkali metal ion or ammonium ion to the pulp where- by the dissolution of desired metals is controlled and the iron content of the solution is reduced to a low level. The leaching may be carried out at temperatures and pressures below or above atmospheric pressure and the atmospheric boiling point of the pulp. (130) Weston, David, 1974b, Hydrometallurgical Treatment of Nickel, Cobalt , and Copper Containing Materials , U.S. Patent 3,798,304, granted March 19. A process for the recovery by leaching of nickel, cobalt, and copper using sulfuric acid and employing agents capable of introducing alkali metal ion or ammonium ion to the pulp where- ^by the dissolution of desired metals is controlled and the iron content of the solution is reduced to a desired level. The leaching may be carried out at temperatures and pressures below or above atmospheric pressure and the atmospheric boiling point of the pulp. (131) Weston, David, 1975, Metallurgical Treatment of Later itic Mineral s, French Patent 2,261,341, granted Sept. 12. Cu, Ni or Co are leached by acid from finely ground lateritic minerals of marine nodules vjith the addition of salts to decrease the solvency of Fe. The ore is ground to a maximum particle size of 590 y with 90% of <_44 y, and made into slurries containing 25% solids (no dispersant) or 60% solids (with dis- persing agent). H2SO4 is added to a pH of < 1.5 (preferably 0.7) and leached at > 50° in the presence of precipitating agents for Fe compounds. These are salts which introduce alkali metal or NH^ ions. The preferred precipitating agents are NaCl, KCl, or their mixtures. A small amount of oxidizing agent (e.g., Cr042-') or Fe203 may be added to improve separation of Fe com- pounds. (132) Wilder, T.C., 1973a, Two State Selective Leaching of Copper and Nickel from Complex Ore , U.S. Patent 3,736,125, granted May 29. Copper and nickel may be selectively leached from a complex ore containing manganese, iron, copper, nickel, cobalt and molybdenum 14-37 by leaching a reduced ore with a solution of an ammonium salt in ammonium hydroxide at approximately room temperature to remove the copper values and then leaching the residue at a higher tempera- ture to extract the nickel. (133) Wilder, Thomas C. and John J. Andreola, 1973b, R ecovery of Copper, Nickel, Cobalt and Molybdenum fro m Comple x Or es, U.S. Patent 3,753,686, granted August 21. Copper, nickel, cobalt and molybdenum may be leached from com- plex ores containing manganese, iron, copper, nickel, molybdenum, and cobalt after the complex ore is roasted with a carbon con- taining material. (134) Wilder, Thomas C. and Walter E. Galin, 1976, Reduction Smelting of Manganese Nodules with a Liquid Reductant , U.S. Patent 3,957,485, granted May 18. A method of recovering metal values from deep sea manganese nodules is presented. A liquid reductant such as bunker-C oil is utilized as a reducing agent. As a result of the efficient contact between the liquid reducing agent and the nodules, little prior preparation of the nodules is required to recovery a high percentage of copper, nickel, cobalt, and molybdenum. A new series of metal alloys are produced in the selective reduction of the nodules, the major phase being iron-rich nickel, one minor phase being 80-90% copper and the other minor phase being a combination of manganese, molybdenum, nickel, cobalt, and iron. 14-38 Cross-Ref erence to Bibliography on Processes 1. Overview Articles 2. Lixiviant: SO2/H2SO4 and other Sulfur Compounds 3. Lixiviant: Chlorides (and other Halides) 4. Lixiviant: Ammonia/ Ammonium Complexes 5. Smelting 6. Metal Recovery 7. Treatment of Red Sea Deposits 8. Mechanisms 9. Pelletizing for Reducing 10. Sewage Sludge 11. Deepsea Ventures 12. Kennecott 13. INCO 14. University of California 15. Bethlehem Steel Ifi. Dow 17. Ocean Resources 18. Bureau of Mines 19. Friedrick Krupp GmbH 20. BAYER 21. Metallgesellschaft AG 22. Other German Work 23. Le Nickel 24. Japanese Work 25. British Patents 26. Canadian Patents 27. French Patents 28. German Patents 29. Japanese Patents 30. South African Patents 31. United States Patents 14-39 1. Overview Articles (usually secondary sources) Agarwal et al, Anon. Barbier Cardwell Cardwell Cardwell Faugeras Granville Hammond Herbert Hubred lammartino Kruger et al. Morgan Pasho et al. Pearson Sisselman Taylor Tinsley Wakefield 1976b 1971 1974 1971 1973a 1973b 1974 1975 1974 1976 1975 1974 1975 1973 1976 1975 1975 1^71 1975 1971 Rev'ew of processes with emphasis on Kennecott's ammoniacal process. Announcement of the development of Deepsea Venture's chloride process. Overview of Processes. Announcement of the development of Deepsea Venture's chloride process. Review of processes with emphasis on Deepsea Venture's chloride process. Review of processes with emphasis on Deepsea Venture's chloride process Overview of Ammoniacal and H2S0^ processes . Nodule characteristics, raining, pro- cessing and economic aspects. Nodule characteristics, origin, mining, processing, and environmental effects. Overview of 1975 activities in nodules processing. Overview with emphasis on characteristics, extractive metallurgy and mining. Processing overview. Overview of processing steps. A comparison of processes. Estimates of nodule reserves are exaggerated . Overview of processes particularly three Deepsea Venture patents. Legal, mining, economics, processes. Announcement of the development of Deepsea Venture chloride process. Processing overview. Announcement of the development of Deepsea Venture's chloride process. 14-40 2. Lixiviant: SO2/H2SO/ and other Sulfur Compounds Brooks et al . 1969 Brooks et al . 1970 Brooks et al. 1971 Faugeras et al. 1975 Fuerstenau et al . 1973 Han et al. 1973 Han et al. 1975a Han et al. 1976b Hanig et al'. 1973 Hanig et al. 1974 Ito et al. 1975 Kane et al . 1974a Kane et al. 1974b Kane et al. 1974c Kane et al. 1974d Kane et al. 1975a Kane et al . 1975d Preliminary study of H2SO4 as leaching agent SO2 leach process for the siliceous nodules . SO2 leach process for the siliceous nodules. Aqueous H2SO4 and SO2 process, recovering Cu/Ni/Co/Mn. H2S0a leach, recovery of Ni/Cu good but slow, Co harder and depends on Fe content Rate data presented. Models for Cu/Ni leaching. H2SO4 leach, Ni/Cu/Co recovery with temperature and oxygen pressure varied. H2SO4 leach, Cu/Co diffusion limited, Ni hetergeneous reaction. Limited study of sensitivity of pH, particle size, dissolution times and temperature. Grain size and time unimportant, tem- perature and surplus acid important. H2SO4 process study. H2SO4 aqueous leach study. Two stage H2Sn4 leach 1) Cu/Ni leach 2) LIX to get Co and Mn 3) Fe++ leach. Roasting Mn nodules with iron sulfide ore in presence of O2 forming iron oxide and sulfates, leaching with water. Selective roast with SO2 without O2 to convert Mn only to soluble oxide and then re-react residue to separate remaining ores. Canadian coverage for 1974b and 1974c plus a SO2 gaseous and a SO2 aqueous processes (same as 1975a?) Aqueous SO2 leach process with LIX tcj separate metals including Mn . H2SO4 leach to obtain Cu/Ni followed, by reducing leach, e.g., ferrous sa its to solubilize Mn/Co. The use of wa ste products as lixiviants pointed out . 14-41 Kane et al 1976a Maekito et al. 1975 Menz 1975 Mero 1965 Meyer-Galow et al . 1973 Neuschutz et al, Weston Weston Weston 1976a Okabe et al. 1974b Schwartz et al. 1974 Ulrich et al. 1973 Van Heck et al . 1973 1974a 1974b 1975 H2SO4 leach to obtain Cu/Ni followed by reducing leach (FeS04) to obtain Mn/Co. H2SO4 aqueous leach study. S02/sulfurous acid leach at 90-100''C. SO2 leach of Ni/Cu/Mn from Co/Fe. SO2 roast and direct aqueous SO2 leach study. H2SO4 high temperature leach followed by extraction step to separate the metals . H2SO4 aqueous leach at 60°C with methanol added . SO2 roast and direct aqueous SO2 leach study. H2SO4 leach. Kinetic study of SO2 absorption and sulfation of Mn nodules. Leaching of Ni from laterite - like ores, using H2SO4 with NH4 present and also the recovery of Ni and Co. Same as 1974a. H2SO4 aqueous leach in presence of pre- cipitating agents such as NaCl or KCl to restrict Fe solubilization. 14-42 3. Lixiviant: Chlorides (and other Halides) Cardwell et al. Cardwell et al. Cardwell et al. Cardwell et al. Cardwell et al. Gandon et al. Hoover et al. Kane et al. Kane et al. Kane et al. Kane et al . Kane et al. Kane et al. Kane et al. 1973a 1973b 1974 1976a 1976c 1976 1975 1973a 1973b 1974c 1975b 1975c 1975d 1976a Meixner et al. 1974 McCutcheon et al. 1975 Queneau et al. 1975 Sandberg et al. 1975 Scheffler et al. 1975 Ulrich et al. 1973 HCl process. Mn/Cu/Ni/Co/Fe203 recovery. Review of chloride process. HCl process. HCl process. Mn/Cu/Ni/Co/Fe203 recovery. HCl aqueous process Mn/Cu/Ni/Co/Fe203 recovery. Smelt at 1700°C followed by HCl aqueous leach recovery Cu/Ni/Co. Various chlorides used: CuCl gave best results at temperatures SOO-IOOO'C. HCl complete process including LIX separation. HCl process with metal chloride vaporiza- tion and condensing. Selective roasting with acid halides to solubilized Mn oxide only. Reaction in molten alkali salt bath to form halides which are vaporized, con- densed in water and separated. Formation of vaporized chlorides which are distilled into fractions and then further processed. HCl leach to obtain Cu/Ni followed by reducing leach (e.g., ferrous salts) to obtain Mn/Co. The use of waste products as lixiviants pointed out. HCl leach to obtain Cu/Ni followed by reducing leach (FeS04) to obtain Mn/Co. HCl aqueous leach study. Halide process in which FeCl3 is recycled. Leach with CI2 in NH4 CI solution after a high temperature reduction. Leaching reduced nodules with metal chloride salts. HCl-Cl2-steam-C02-N2 mixture was used to chlorinate nodules. HCl leach. 14-43 4. Lixiviant : AmTnonia/Ammonium Complexes Anglo-American Corp. 1976 Bare et al . Earner et al. Rarner et al. Brooke et al. Brooks et al. Han et al. Okabe et al. Oka j ima Pemsler et al, Pemsler et al. Redmond Rolf Skarbo Skarbo Szabo Szabo Szabo Wilder Wilder 1973 1976a 1976b 1969 1970 1974 1974a 1975 1976a 1976b 1973 1969 1973a 1973b 1975a 1975b 1976 197 3a 1973b Leaching of Ni/Co oxides with mixture of ammonium sulphite and bisulphite. Ni/Cu/Co recovery. Ammoniacal Cuprion process - use Cu at >2g/l for efficiency. Fluidized bed drying and reduction of nodules followed by NH3-(NH4) .CO-j leaching. A few tests performed. Ammoniacal leach process for calcareous nodules . Ammonium carbonate/sulfate leaching after CO/CO2 reduction. Ammonium sulfate leach. Ammonium carbonate leach with NaCl added after reduction roast. Na2S and Na„SO- used to recover metals. Cuprion process followed by LIX64 extrac- tion of slurry. Solvent-in-pulp extractions of Cu and Ni from ammoniacal leach slurries. CO2/CO reduction followed by NH3/CO2 leach. Ammonium sulfate leach. Ammoniacal leach. Ammoniacal leach at high T and P. Ammoniacal cuprous ion leach and reduc- tion. Cuprion process is improved by maintaining Cu concentration >10 g/1. Basic U.S. patent on Cuprion process. Ammoniacal salts in NH^OH solution; Cu/Ni leach after reduction. Ammoniacal salts in NH4OH solution; Cu/Ni leach after carbon reduction. 14-44 5 . Smelting Beck et al. 1971a Beck et al. 1971b Bell et al. 1975a Bell et al. 1975b Bell et al. 1976 Gandon et al. 1976 nils et al. 1975 Nagaya et al. 197 3 ■ Sridhar 1974 Sridhar 1975 Sridhar et al. 1976 Vasllchikov et al. 1968 Wilder et al . 1976 With petroleum coke as reductant. Smelting with sulfur and carbonaceous reducing agent added. Recovery of metals from matte by fused salt method. Smelting at >1100°C followed by mag- netic separation of ground matte. Smelt at >1100°C to reduce the metal values and coalesce into discrete metallic particles. Smelt at 1700°C followed by HCl aqueous leach to recover Ni/Cu/Co. Smelt at 1200°C followed by NH3/CO2 leach for recovery of Ni/Cu/Co. Residual treated with SO2 solution to recover Mn. Smelt at 1500°C after reduction at 1000°C Mn slag smelted again to form ferro- manganese. With reducing gas CO/H2 1100-1259°C. Smelt at 1250-1600°C after reduction at 900°C. The matte sulfided and processed by conventional means to obtain metal values . Thermal reduction followed by H2S0^ pressure leach. Thermal reduction of nodules. Reduction smelting of manganese nodules with a liquid reductant. 14-45 6. Metal Recovery Agarwal et al. 1976a Barton et al . 1974 Bell et al. 1975a Cardwell et al . 1974 Cardwell et al . lQ75a Cardwell et al. 1975b Cardwell et al . 1976a Cardwell et al . 1976b Coltrinari 1972 Faugeras et al. 1975 Illis et al. 1975 INCO 1975a TNCO 1975b Kane et al. 1973a Kane et al. 1974e Kane et al . 1976b Kauczor et al. 1974 McCutcheon et al. 1974 Pemsler et al. 1976a Pemsler et al. 1976b Cu/Ni separation using LIX64 fluid ion exchange. Mn from Mn halides by fused salt electro- lysis . Recovery of metals from matte by fused salt method. Liquid ion exchange to separate Cu/Co/ Ni/Mn. Ni/Co separation by liquid ion exchange. Ni/Co separation by liquid ion exchange. Mn/Fe/Ni/Cu/Co halide separation by liquid ion. Mn from Mn halide by reaction with Al . Separation Ni from Co in NH3 solutions using LIX64 . Recovery of Cu/Ni/Co/Mn using LIX and solvent extraction after H2SO4 aqueous leach. Smelt at 1200°C followed by NH3/CO2 leach for recovery of Ni/Cu/Co . Residual treated with SO2 solution to recover Mn. LIX of Cu, Ni, Co, or Zn from solution containing Si02. LIX of NiS04 with Co, Cu and Zn present. LIX to separate Cu/Ni/Co chlorides from Mn chlorides. Mn/Fe separation: convert to halides and using Al to obtain Mn. Metal halides reacted with Fe/Mn metals form Fe/Mn halides and pure metals (Cu/Ni/Co). LIX for all metals except Mn using Lewatit TP207 and selectively removing metals . Improved electrolytic recovery of metals (Ni, Co, Sn, Fe and Mn) from metal halides. Cuprion process followed by LIX64 extrac- tion of slurry. Solvent in pulp extraction of Cu/Ni from ammoniacal leach slurries. 14-46 Roever et al. Schoebert et al. Skarbo Skarbo Skarbo Skarbo Skarbo et al. Skarbo et al. Ska;rbo Skarbo Skarbo et al. Skarbo et al. 1975 1976 1974a 1974b 1974c 1974e 1974f 1974g 1975a 1975b 1976a 1976b Use of ion exchange resin to separate Cu from Co/Ni/Fe/Mn solution. Convert MnO to Mn via metal halide. Ni from metal salts Cu/Ni, Co and impurities . Cu from metal salts Cu, Ni, Co and impurities. Selective stripping from oximes. Recovery of Ni from Cu by separate stripping from an oxime extractant . Removal of NH3 from loaded oxime extractant. Removal of NH3 from loaded oxime extractant. Cobalt stripping from oximes. Ni extraction and stripping using oximes and ammoniacal carbonate solutions. Process for recovering Cu from basic carbonate mixtures. Separation of Cu from Ni at pH >9. 14-47 7 . Treatment of Red Sea Deposits Clement et al. Halbach Neuschutz et al< Supp et al. 1973 1976 1976b 1973 Cu/Zn sulfides separated by floatation. Reduction of metals using sewage sludge. H2SO4 pressure leach with O2 of Red Sea concentrate - good recovery but un- economic . Concentrating Red Sea ooze for further processing . 8. Mechanisms Burns Burns et al , 1965 1966 Cobalt is in amorphus iron phase of sea nodules. Cobalt in iron phase; Ni/Cu in Mn phase- 9 . Pelletizing for Reducing Agarwal et al. Moreau et al. 1974 1976 Method of pelletizing. Manganese nodule pelletizing, 10. Sewage Sludge Halbach et al. 1976 Sewage sludge as reductant in treatment of deep sea nodules. lA-48 11. D eep Sea Ventures I Barton et al. Cardwell Cardwell Cardwell et al, Cardwell et al . Cardwell et al . Cardwell et al, Cardwell et al. Cardwell et al. Coltrlnari Kane et al . Kane et al. Kane et al . Kane et al. Kane et al. Kane et al. Kane et al. Kane et al. Kane et al. Kane et al. Kane et al. 1976 1973a 1973b 1974 1975a 1975b 1976a 1976b 1976c 1972 1973a 1973b 1974a 1974b 1974c 1974d 1974e 1975a 1975b 1975c 1975d Mn from fused salt electrolysis. Overview of processes. Emphasis on chloride process. Overview of processes. Emphasis on chloride process. HCl process: recovers Mn/Cu/Ni/Co/ Ni/Co separation via liquid ion exchange. Ni/Co separation via liquid ion exchange. HCl process recovers Mn/Cu/Ni/Co/Fe. Mn from Mn halide by reacting with Al. HCl aqueous process Mn/Cu/Ni/Co/Fe203 recovery . Separation of Ni from Co in NH3 solutions using LIX64 . HCl process. HCl process with vaporization and con- densing of metal chlorides. Two stage leach with H„SO^. Roasting Mn nodules with iron sulfide ore to form sulfates and iron oxide. Selective roast with acid reagent with- out O2 to solubilize Mn oxides only. Canadian coverage for 1974b and 1974c plus SO2 gaseous and aqueous processes (1975a). Mn/Fe separation by converting to halides and using Al to obtain Mn. Aqueous SO2 leach process followed by LIX to separate metals. Formation of halides in molten salt bath which are then vaporized. Formation of vaporized halides which are distilled into fractions. Acid leach for Cu/Ni followed by reducing leach for Mn/Co. Waste products can be used as lixiviants. 14-49 Kane et al. 1976a Kane et al. 1976b McCutcheon et al . 1975 McCutcheon et al. 1975 Sandberg et al . 1975 Schobert et al. 1976 Acid leach for Cu/Ni followed by reducing leach (FeSO^) for Mn/Co. Metal halides (liquid or vapor) reacted with Fe/Mn metal to form Fe/Mn halide and pure metals (Cu, Ni, Co). Electrolytic recovery of metals from metal halides. Halide process in which FeCl3 is recycled . Leaching reduced nodules with metal chloride salts. Mn recovery from MnO by reaction with subhalide to transport metal, e.g., Al. 14-50 12. Kennecott Agarwal et al. 1974 Agarwal et al. 1976a Agarwal et al. 1976b Earner et al. 1976a Earner et al. 1976b Beck et al. Beck et al. Hubred Moreau et al. Pemsler et al, Pemsler et al, Redman Skarbo Skarbo Skarbo Skarbo Skarbo Skarbo Skarbo Skarbo et al. Skarbo et al. Skarbo et al . Skarbo et al. Skarbo et al. 1971a 1971b 1975 1976 1975a 1976b 1970 1973a 1973b 1974a 1974b 1974c 1974d 1974e 1974f 1974g 1975a 1975b 1976a Pelletizing nodules before reduction. Gu/Ni separation by LIX64 fluid ion exchange . Overview of processes with emphasis on Kennecott 's antmoniacal process. Ammoniacal Cuprion process - use Cu at >2 g/1 for efficiency. Fluidized bed drying and reduction of nodules followed by NH3-(NH4) 2CO3 leaching. Smelting with petroleum coke as reductant. Smelting with sulfur and carbonaceous reducing agent added. Overview with emphasis on nodule charac- teristics, extractive metallurgy and mining . Manganese nodules pelletizing. Cuprion process followed by LIX64 extrac- tion of slurry. Solvent-in-pulp extraction of Cu and Ni from ammoniacal leach slurries. Ammoniacal-of f gas process. Ammoniacal process - Mn reduction/ leach. Ammoniacal process at high T and P. Ni recovery from metal salt mixture. Cu recovery from metal salt mixture. Cobalt stripping from oximes. Selective stripping of nickel from oxime. Recovery of Ni from Cu by separate stripping from an oxime extractant. Removal of NH3 from loaded oxime extractant. Removal of NH3 from loaded oxime extractant. Cobalt stripping from oximes. Ni extraction and stripping using oximes and (NH4)2C03. Process for recovery Cu from basic car- bonate mixture. 14-51 Skarbo et al, Szabo Szabo Szabo Wilder Wilder Wilder et al 1976b 1975a 1975b 1976 1973a 1973b 1976 Separation of Cu from Ni at pH >9. Cuprion ammoniacal process with Cu"^ reducing Mn"^^ and regenerated with carbon monoxide. Cuprion process is improved by maintaining Cu"^ concentration >10g/l. Basic U.S. Patent on cuprion process, Ammoniacal salts in NH4 OH solution Ni/Cu leach. Ammoniacal salts in NH4OH solution Ni/Cu leach after carbon reduction. Reduction smelting with a liquid reductant . 14-52 13 . International Nickel Company Bell et al. 1975a Bell et al. Bell et al. Illis et al, INCO 1975b 1976 1975 1975a INCO 1975 Sridhar 197A Sridhar et al. 1975 Sridhar et al. 1976 Weston 1974a Weston 1974b Weston 1975 Recovery of metals from matte by fused salt method. Smelting at >1100°C followed by mag- netic separation of ground matte. Smelt at >1100°C to reduce the metal values and coalesce into discrete metallic particles. Smelt at 1200°C followed by NH3/CO2 leach for recovery of Ni/Cu/Co. Residual treated with SO2 solution to recover Mn. LIX of Cu, Ni, Co, or Zn from solution containing Si02 • LIX of NiSO^ with Co, Cu and Zn present. Thermal reduction. Smelt at 1250-1600°C after reduction at 900°C. The matte sulfided and processed by conventional means to obtain the metal values. Sulfide formation after pyro reduction. Ni, Cu, Co leached from laterites with H2S0^ with NH4+ present. Ni, Cu, Co leached from laterite-like ores with H2SO4 with NH^+ present . H2SO4 aqueous leach in the presence of precipitating agents such as NaCl or KCl to restrict Fe solubilization. 14-53 14. University of California Han et al . 1973 Han et al. 1974 Han et al. 1975a Han et al. 1976b Hoover et al. 1975 Fuerstenau et al . 1973 H2SO4 leaching model for Cu and Ni. Ammonium carbonate/sulfate leach after CO/CO2 reduction. Ni/Cu/Co leaching with H2SO4. H2SO4 leaching study. Roasting with various chlorides. CaCl2 the best. H2SO4 leaching of Cu, Ni, and Co. 15. Bethelehem Steel Bare et al, 1973 Ammoniacal leaching after reducing with CO/CO2. 16. Dow Rolf 1969 Ammonium sulfate leach process, 17. Ocean Resources Mero 1965 SO2 leach separating Ni/Cu/Mn from Co/Fe. 18. Bureau of Mines Brooks et al. Brooks et al . 1970 1971 Ammoniacal and SO2 leach processes. Ammoniacal and SO2 leach processes. 19 . Friedrick Krupp GmbH (Germany) Meixmer et al. Neuschutz et al, Scheffler et al . Supp et al. Ulrich et al. 1974 1976a 1974 1973 1973 HCl aqueous leach study. H2SO4 high temperature leach of Mn nodules . HCl-Cl2-Steam-C02-N2 mixture was used to chlorinate the nodules. Concentrating Red Sea ooze by floata- tion methods. HCl and H2SO4 leach. 14-54 20. Bayer (Germany) Kauczor et al , Roever et al. 1973 1975 LIX separation of metals from Mn. Use of ion exchange resins to separate Cu from Co/Nl/Fe/Mn solution. 21 . Metallgesellschaf t AG (Germany) Kruger et al . 1975 Meyer-Galow et al. 1974 Schwarz et al. 1974 Overview of processing steps. S0« roast and aqueous SO2 leach study, SO2 roast and aqueous SO2 leach study. 22. Other German Work Anon. 1976 Clement et al. 1973 Halbach et al. 1976 Hanig 1973 Hanig et al . 1974 Kruger 1973 Menz 1975 Interocean '76. Cu/Zn sulfides separated from Red Sea sediment . Metal reduction using sewage sludge. H2SO, leaching of Cu/Ni/Co. H2SO4 process study. Interocean '73. S02/sulfurous acid leach at 90-100°C, 23. Le Nickel Gandon 1976 Smelt at 1700°C followed HCl aqueous leach to recover Cu/Ni/Co. 24 . Japanese Work Ito 1975 Maekito et al. 1972 Nagaya et al. 1973 Okabe et al . Okabe et al. Oka j ima 1974a 1974b 1975 H2SO4 aqueous leach. H2SO4 aqueous leach. Smelt at 1500°C after reduction at 800°C. Mn slag smelted again to form ferro- manganese. Ammonium sulfate leach. H2SO4 leach at 60°C with MeOH added. Ammonium carbonate leach with NaCl added after a reduction roast. Na2S and Na2S03 used to recover metals. 14-55 25. British Patents Number Date Author 1,425,845 1,423,584 1,445,123 02/18/76 02/14/76 08/04/76 Cardwell et al. Anglo American Corp. Skarbo 26. Canadian Patents 871,066 951,908 974,371 989,179 965,960 05/18/71 07/30/74 09/16/75 05/18/76 04/15/76 Beck et al. Kane et al. Illis et al. Bell and Sridhar Roever et al. 27. French Patents Number 2,211,453 2,236,949 2,244,004 2,261,341 2,264,876 Date Author Skarbo Concordance 07/19/74 German 2,360,147 02/07/75 Bell 04/11/75 Inco 09/12/75 Weston 10/17/75 Inco 14-56 28, German Patents Number Date Author Concordance 2,126,175 /70 Kane et al. U.S. 3,854,851 2,126,222 01/05/72 Kane et al. U.S. 3,809,624 2,135,733 01/20/72 Wilder U.S. 3,753,6J?6 2,152,696 111 Coltrinari 2,156,940 05/31/74 Skarbo U.S. 3,728,105 2,234,971 {^llllllk Queneau et al. U.S. 3,880,651 2,240,586 oimiih Maxiner, Scheffler, and Ulrich 2,247,497 04/05/74 Agarwal et al. U.S. 3,788,841 2,249,302 04/11/74 Scheffler and Ulrich 2,332,959 01/10/74 Skarbo Brit. 1,445,173 2,332,976 01/10/74 Skarbo U.S. 3,853,725 2,333,003 01/17/74 Skarbo et al. U.S. 3,855,090 2,359,642 06/06/74 Szabo South African 8885/73 2,360,147 01/27/74 Skarbo 2,364,550 05/09/74 British 1,425,845 2,438,773 03/20/75 Bell and Sridhar 2,507,930 09/04/75 Faugeras et al. 2,510,243 09/20/75 Sridhar 2,522,280 12/30/75 Schobert et al. U.S. 3,950,162 2,522,954 12/30/75 Sandberg et al. 2,522,969 11/27/75 Oka j ima 2,524,598 12/11/75 Szabo 2,526,388 01/02/76 Barner et al. 2,526,395 01/02/76 Pemsler et al. 2,530,909 01/22/76 Gandon 2,537,940 03/11/7 6 Barner et al. 14-57 29. Japanese Patents Number 72 30,505 73 56,513 74 123,110 74 123,111 Date 11/09/72 08/08/73 11/25/74 11/25/74 Author Maekita et al. Nagaya et al. Okabe et al. Okabe et al. 30. South African Patents Number 8885/73 Date 02/97/75 Author Szabo Concordance U.S. 3,983,017 14-58 3; L. I Numl J.S. Patents Der 3 ,169 ,836 3 ,471 285 3 ,723 095 3 ,728 105 3 ,734 715 3 ,736 125 3 ,751 554 3 ,752 745 3 ,753 '686 3 ,773 635 3 ,788 841 3 ,793 430 3 795 596 3 798, 304 3 798, 307 3 ,809, 624 3 810, 827 3 810 971 3 ,832, 165 3 832, 295 3 849 534 3 853, 725 3 ,854, 851 3 855, 089 3 ,855, 090 3 ,867, 506 3 ,869 360 3 ,880, 651 3 ,894 139 3 894 927 Date 02/16/65 10/07/69 03/27/73 04/17/73 05/22/73 05/29/73 08/07/73 08/14/73 08/21/73 11/20/73 01/29/74 02/19/74 03/05/74 03/19/74 03/19/74 05/07/74 05/14/74 05/14/74 08/27/74 08/27/74 11/19/74 12/10/74 12/17/74 12/17/74 12/17/74 02/18/75 03/04/75 04/29/75 07/08/75 07/15/75 Author Mero Rolf Skarbo Skarbo Redmond et al. Wilder Bare et al. Kane et al. Wilder et al. Kane et al. Agarwal et al. Weston et al. Kane et al. Weston et al. Skarbo Kane et al. Kane et al. Skarbo et al. Kane et al. Barton et al. Skarbo Skarbo Cardwell et al. McCutcheon et al. Skarbo Skarbo et al. Kane et al. Queneau et al. Cardwell et al. Kane et al. 14-59 31. U.S. Patents (continued) Number 3,901,775 3,903,235 3,903,236 3,906,075 3,907,966 3,923,615 3,930,974 3,940,470 3,942,974 3,950,162 3,950,486 3,950,487 3,950,488 3,951,647 3,957,485 3,983,017 Date Author 08/26/75 09/02/75 09/02/75 09/16/75 09/23/75 12/02/75 01/06/76 02/24/76 03/09/76 04/13/76 04/13/76 04/13/76 04/13/76 04/20/76 05/18/76 09/28/76 Kane et al. Cardwell et al. McCutchern et al. Menz Skarbo Kane et al. Kane et al. Kane et al. Moreau et al. Schobert et al. Cardwell et al. Pemsler et al. Skarbo et al. Cardwell et al. Wilder et al. Szabo 14-60 *U.S. GOVERNMENT PRINTING OFFICE : 1977 0-248-057/6614 K I I u ^ \ I PENN STATE UNIVERSITY LIBRARIES illilllllliilllllli A0DDD7EQEaST7 NOAA--S/T 77-2991 Vol. Ill It-